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By Miro Mihelich

The mines and prospects in the vicinity of Landlocked Bay, 25 miles southwest of Valdez, Alaska (fig. 1), were examined, mapped, and sampled in July and August 1955 as part of a comprehensive investigation by the Bureau of Mines of ore deposits adjacent to Prince William Sound. The purpose of the investigation is to establish and record the location, nature, extent, and metallurgical characteristics of copper and associated metallization in the Prince William Sound district and to indicate, if possible, those areas in which additional exploration and development may be justified. Of particular interest are the possibly numerous deposits of massive iron sulfides that were neglected during early mining operations. These deposits may constitute important reserves of sulfur and iron. At Landlocked Bay deposits of copper-bearing iron sulfide minerals have been mined or partly explored through mine workings, which are in a zone about 1 mile wide along the western side of a large overthrust fault known as the Landlock Overthrust. Eight of the principal mines along this zone were examined end sampled. Although only three of the mines have a recorded production of shipping-grade copper ore, all have deposits of copper-bearing iron sulfide minerals that are similar in occurrence. The deposits occur as lenticular masses and disseminations of chalcopyrite, pyrite, and pyrrhotite which replace crushed rock or fill cavities along shear zones in greenstones and in sedimentary rocks that are exposed between sea level and an altitude of 2,500 feet. These lenticular masses of sulfide minerals range in size from small pods to ore bodies that have been stoped to maximum widths of 9 feet, strike lengths up to 70 feet, and dip lengths as much as 350 feet. Where accessible for sampling,, the copper content of the deposits ranged from a few hundredths of a percent to over S percent; zinc, gold, and silver are present in negligible amounts. Flotation tests of a sample composited to represent typical mine-run ore indicate that 81 percent of the copper is recoverable in a product assaying 28 percent copper at a concentration ratio of 17.5 : 1. INTRODUCTION Statistical records indicate that approximately 110,000 tons of metallic copper (about 16 percent of total Alaskan production) was produced from mines adjacent to Prince William Sound during the 40-year period ended in 1938. Although the principal production was from a large-scale mining and milling operation on Latouche Island (Kennecott Copper Corp., Beatson mine), a substantial amount of the total copper was derived from direct-shipping ores selectively mined or hand-sorted from numerous other deposits in the district. No production has been made since 1938; many of the claims have long been abandoned, and mine workings are either caved or deteriorating rapidly.

Jan 1, 1957

By Walter P. Jenkey

[ ] I. THE REDUCING ACTION OF CARBON AND OF HYDROCARBONS. Carbon has long been recognized as one of the most powerful reducing agents in the deposition of ores. Investigations, made by myself, of the zinc- and lead-deposits in Southwest Missouri, in the region centering about Joplin, where the formation of the metallic sulphides has been due to the action of bitumen, carbonaceous shales and bituminous coal, Lace afforded abundant evidence that the solid oxygenated hydrocarbons, particularly, when in fine powder and in suspension in the waters circulating through the ore-bodies, are the most energetic and powerful reducing agents known. Bitumen, liberated by the decomposition of the ore-bearing limestone, is found in the Joplin mines in all degrees of fluidity and hardness, dependent on the amount of oxidation it has undergone. From semi-fluid maltha it grades into partly oxidized mineral-pitch, which, by further oxidation, changes to hard asphalt, finally breaking up, from continued absorption of oxygen, into a fine powder resembling in appearance powdered coal. In this condition oxidized bitumen, from its light gravity, is transported readily in suspension in the underground circulating-waters. Bituminous coal and black carbonaceous clays and shales occur as surface formations, often in intimate association with

Jan 1, 1913

By S. Carsten, S. Subramanian, R. Vasant Kumar, S. Vimalnath

"The effect of EPS components namely protein, polysaccharide and biosurfactant purified from Pseudomonas fluorescens was investigated for the bioremediation of Pb(II) ions in aqueous solution. Two parameters, namely the amount of EPS component and the concentration of Pb(II) ions were optimized for maximum Pb(II) ions uptake. The EPS components interacted with Pb(II) ions were characterized using energy dispersive X-ray analysis (EDAX) to ascertain the presence of Pb and by attenuated total reflection-Fourier transform infrared spectroscopy (ATR–FTIR) to reveal the functional groups involved in Pb(II) binding. The maximum Pb(II) binding capacities of protein, polysaccharide and biosurfactant were found to be 11.4, 0.55 and 0.11 mg mg-1 respectively. The biosorption processes followed the Langmuir model. The Gibbs free energy (?G0) values of -23.6, -20.9 and -23.2 kJ mol-1 for Pb(II) biosorption by protein, polysaccharide and biosurfactant respectively, attested to spontaneous, chemisorption processes. The ATR–FTIR results revealed the presence and involvement of amino, hydroxyl, carboxyl, phosphate, and methyl functional groups in Pb(II) binding and EDAX analysis confirmed the interaction of Pb with EPS components. The Pb(II) binding mechanisms of EPS have been discussed. The results demonstrate the potential of the purified EPS components of P. fluorescens for removal of Pb(II) ions from aqueous solution INTRODUCTION Waste water containing toxic metals from various industries pose a serious threat to human health and ecosystem. For treating such waste water, conventional methods like chemical precipitation, coagulation, evaporative recovery, electrodialysis, floatation, flocculation, ion exchange, reverse osmosis, nanofiltration, ultrafiltration are used (Lakherwal et al., 2014). These methods have some disadvantages like toxic sludge generation, incomplete metal removal and high cost (Singh and Gadi, 2012), necessitating the search for cheaper and efficient alternative methods. Bioremediation is one of the emerging alternatives where bacteria, fungi, yeast, algae and other biological substances are used as sorbents to remove toxic heavy metals (Vijayaraghavan, et al., 2015). Recently, the bioremediation potential of extracellular polymeric substances (EPS) of microbial origin has attracted attention. However, for Pb(II) removal, there is a paucity of literature focusing on EPS."

Jan 1, 2018

By N. J. Grant, B. C. Giessen, M. Morris

The slvuclures of alloys in the system In-Bi have been investigated after (levy vapid queuching from the mell (splat cooling) to -190°C. Tuo-phase fields could be suppressed over most of the tota1 concentvalion range; five melastable phases a1, a2, ?, ?1. and ß exisl, of which four have simple elementtike structures. For example, ß has the A5 struclure of white till. Crystal-tographic data and coordination numbers for these phases aye given, as well as addilional information on the equilibrium diagram, where a new phase In5Bi3 teas found. THE phase diagrams of binary combinations of B elements from the groups B2, B3, B4, and B5 are usually of a simple type and show considerably fewer intermediate phases than combinations of these elements with A metals, transition metals, or the B1 noble metals. Aside from phases such as the typical B3-B5 compounds with the ZnS-B3 structure type, few stoichiometric phases are known; among them are In2Bi and 1nBi.l Nonstoichiometric phases include ß (In-Sn), a1, (In-Pb), and the tin-rich ? phases in the In-Sn, Cd-Sn, and Hg-Sn systems.' The first two phases have a close relationship to indium, while the latter group are known to be structurally related to a Sn. Thus, there are generally no immediate experimental data for the correlation of the crystal structures of B metals and B metal combinations with certain monotonously varying parameters, such as the valence electron concentration (VEC), or the average atomic size. Such parameters have been recognized as structure determining, e.g., for the transition 4 Sn — y phase,2,3 where nonintegral valence electron concentrations are encountered. It should be possible to extend such correlations to other binary systems, if nonequilibrium single-phase alloys could be produced over broad valence electron concentration ranges, and if their crystal structures could be regarded as being imposed by their electronic states. It has been shown in the case of the tin-rich 7 phases3 that alloy phases with typical crystal structures can be produced in B metal alloys rapidly quenched from the liquid to -190°C by the splat cooling technique due to Duwez,3-8 and reviewed in Refs. 5 and 6. The In-Bi system was selected because it extends over a valence electron concentration range in which several metastable, nonstoichiometrir phases could be expected to occur. Further, the low melting points of the known intermediate phases, In2Bi and InBi, Fig. 1(a), indicated low binding energies and thus possibly low driving forces for the formation Of the equilibrium structures. EXPERIMENTAL TECHNIQUES AND RESULTS The preparation of the quenched foils followed the practice described in Ref. 3. Master alloys were produced by melting of indium (99.97+) and bismuth (99.99+) in evacuated Vycor capsules or in an inert gas arc furnace; quantities of 20 mg were splat-cooled onto copper and silver substrates held at - 190°C; and crystal structures and lattice parameters were determined on a GE XRD-5 diffractometer using Cu Ka, radiation at -190°C and at room temperature. The duplication of each run with both substrates permitted the elimination of overlap of the substrate diffraction pattern and that of the investigated substance. The XRD patterns were usually taken from sin2 " = 0.03 to 0.35; the substrate was used as a means of internal calibration. The fractional accuracy of the lattice parameters of metastable phases is approximately 10-3 In the following, all percentages are in atomic percent. The stable and metastable phases found after rapid quenching to -190°C are listed in Table I, together with estimated concentration ranges and crystallographic data. The investigated alloys and phases present in them are given in Table 11. Except for the region between indium with 33 and 50 pct Bi, where a revision of the equilibrium phase diagram became necessary to include a new stoichiometric phase In5Bi3, the diffraction patterns taken after heating to room temperature agreed with those expected from the phase diagram, Fig. l(a). This suggests that the new nonstoichiometric phases are not stable at room temperature, thus following the observations made for the y phases based on tin.3 Results concerning the equilibrium phases In2Bi and In5Bi3 and the revision of the equilibrium phase diagram which is made necessary by the inclusion of In5Bi3 will be treated first. The Crystal Structure of In2Bi. Makarov7 had identified In2Bi as belonging to the AlB2-C32 type ; however, in a later paper the structure was revised.8 In2Bio was found to be of the Ni21n-B8ß type,21 with a = 5.496A, c = 6.57.9A, and N = 6 atoms per unit cell.9 This crystal structure was confirmed in the present work; definite evidence for the doubling of the c axis, as compared to the A1B2-C32 type, and for the proposed order structure, was found in the powder pattern. The observed lattice parameters agree well with those of Ref. 8; those measured at -190°C are given- in Table I. The Crystal Structure of In5Bi3. A new, equilibrium, intermediate phase was identified in slow-cooled, as-cast alloys; it was identified as In5Bi3. The structure has been worked out by Giessen and Grant;" lattice parameters at -190°C are given in Table I. This phase is probably identical with "In3Bi2", produced by vapor deposition techniques, but not recognized as an equilibrium phase by Palatnik et a1.11 After completion of the present work, the existence of In5Bi3 has also been demonstrated by superconductivity measurements on In-Bi alloys.20

Jan 1, 1968

By M. J. Ross, D. A. Norrgran, J. A. Wernham, J. N. Orlich

INTRODUCTION Laboratory testing for magnetic concentration circuit design is a requirement, just as it is for most other types of concentration. Natural variations in raw materials allow on1 y the broadest of generalities to be made regarding magnetic separation. Further, magnetic fields are geometry dependent, leaving fundamental models of a separator overly dependent upon the exact conditions for which the models were derived. Despite this, an understanding of the basics of magnetic separation and of the problem to be attacked can result in rapid progress towards the final circuit design with a relatively simple testing program. Magnetic Separation Basic Principles A magnetic field can readily be understood in terms of lines of flux. A familiar demonstration of these lines of flux is the pattern of iron filings on a sheet of paper when a magnet is placed underneath. Lines of flux2 are measured in Gauss (1 line/cm2) or Tesla (10000 gauss). Regions of high field strength have a greater number of lines. Field direction is indicated by a convergence of lines. When a particle is placed in a magnetic field it cuts the lines of flux and alters them in one of three basic patterns. These patterns are referred to as diamagnetic, paramagnetic and ferromagnetic. Diamagnetic materials have electronic structures which react to oppose an applied magnetic field. This opposition acts to locally reduce the applied field slightly. The visualization is that lines of flux diverge and go around the particle. Because of this divergence or opposition, a diamagnetic particle will tend to move towards regions of lower flux density. That is to say that the particle will be repelled by a magnetic field. Aluminum oxide and napthalene are examples of diamagnetic materials. Paramagnetic materials have an electronic structure which increases an applied magnetic field. Although these materials also exhibit diamagnetism, they have randomly oriented impaired electrons which align with the magnetic field. The resultant local field enhancement is significantly larger than the diamagnetic opposition. The visual-

Jan 1, 1986

By S. R. Maloof, W. R. Mitchell, J. A. Mullendore

Preliminary experimental evidence is presented to show that the metallic impurities aluminum, iron, and silicon, and beryllium oxide as found in commercially pure hot-pressed beryllium powder can be reduced to lower Concentrations by Zone -Purification techniques. The reduction in the concentration of aluminum to extremely low levels (10 PPm) is noteworthy, since earlier work demonstrated that aluminum is a major factor contributing to the hot-tearing of beryllium during fusion welding. On the basis of present findings, a method is suggested for producing beryllium metal of improved weldability. EXPERIMENTAL research now being conducted on the purification of beryllium indicates that several of the metallic impurities and beryllium oxide can be reduced to lower concentrations by employing standard horizontal zone-refining techniques. For the preliminary work, commercially-pure hot-pressed beryllium having a purity of 97.84 pct was used. The chemical composition is given in Table I. The hot-pressed beryllium was machined into bars 5/8 by 5/8 by 4 to 6 in. in length. Beryllium-oxide boats which had been conditioned by prior heating in a vacuum of 10"5 mm of Hg at 1100°C for 3 hr were used to hold the beryllium while it was zone melted. The boat with its charge, along with a smaller beryllium-oxide boat containing scrap beryllium to be used for gettering purposes, was loaded into a quartz tube and evacuated to 10-6 mm of Hg. Alter alternately evacuating and flushing three times with purified argon, the furnace tube was backfilled to a pressure of 1/6 atm of A and sealed off by means of a vacuum-tight stopcock. It was then removed from the pumping equipment and mounted on the traveling carriage of the zone-refining apparatus. Heating was accomplished with a 10 kw 450 kc induc- tion heating unit using an induction coil consisting of two turns of 3/16-in. diam copper tubing. Before starting to zone melt, the atmosphere in the furnace tube was gettered by heating the scrap beryllium in the smaller boat to above its melting point and holding at temperature for several minutes. The carriage was then shifted so as to locate one end of the beryllium ingot in the large boat within the induction coil. A molten zone approximately 3/4 to 1 in. was established by loading directly into the ingot. The molten zone was moved very slowly along the length of the ingot. Before each pass was made the procedure for evacuating, flushing, and backfilling, and so forth was repeatedI' Speeds of 1 1/2 and 3 in- per hr were used with up to twelve zone passes. The appearance of a typical zone-refined ingot is shown in Fig. 1. After zone refining, the grains in the ingot appear to be elongated in the direction of zone travel, whilst before, they are essentially equi-axed. A microexamination revealed the presence of a network in the beryllium grains as seen in Fig. 2. Laue spots in back reflection pictures taken of the sample were not sharp, indicating the presence of subgrains in the structure. As can be seen, however, many of the network boundaries cross beryllium grain boundaries. It has been shown by Martin and Moore1 that beryllium solidifies as a bcc structure which then transforms to the hexagonal structure at about 1250oC.It is believed that the observed network is merely the outline of the bcc structure which formed from the melt and are not subgrain boundaries. In order to eliminate surface contaminants, approximately 1/16 of an inch of material was removed from the outer skin of each ingot after zone refining.

Jan 1, 1962

By M. E. Volin

The Buckeye manganese deposits are in Stanislaus County, Calif., about 33 miles south of Tracy. From 1917 to 1919 the deposits produced 13,370 tons of ore reported to average 46 percent manganese and 14 percent silica. The deposits are confined to two areas of' development, known as the north mine workings and the south mine workings, about 700 feet apart horizontally. Mineralization occurs chiefly as rhodochrosite and psilomelane interbedded in cherts of the Franciscan formation. As a part of the Bureau of Mines program of investigation and development of domestic mineral deposits, a development project was started November 25, 1940, and completed June 27, 1941. The Federal Geological Survey cooperated by mapping the topography and geology and logging the drill cores. The program included road and. trail building, trenching, underground excavation, diamond drilling, and sampling. The following pages mention the location, brief the history, and note the physical .features of' the district. Mine workings are sketched, the deposit is described, and the plan of the project is noted. The work performed by the Bureau of Mines is explained in the concluding pages with a summary of the drill-hole logs.

Jan 1, 1949

By Leopoldo José Naves Alves

Os sistemas de controle e supervisão de processos que não contemplam o adequado gerenciamento de alarmes e eventos sofrem uma sobrecarga na operação com um número exacerbado de informações durante situações de distúrbio e até mesmo operação normal. Nos últimos anos, a abordagem sobre metodologia, boas práticas e planejamento adequado do gerenciamento de alarmes vem sendo alvo de crescente interesse, seu objetivo é garantir um ótimo estado de produção e uma maior segurança dos equipamentos e dos operadores. Em 2009 a ISA, após anos de trabalho sobre este tema, apresentou a norma ANSI/ISA 18.2, formalizando conceitos, definindo metodologias e sugerindo boas práticas. Este trabalho visa apresentar de maneira objetiva tal norma e demonstrar a viabilidade e eficiência de sua aplicação pratica.

Aug 8, 2018

By Eckley B. Coxe

Having been appointed, on February 19th, 1890, a member of the Commission created by the Legislature of Pennsylvania for the purpose of investigating the " Waste of Coal Mining, with the View to the Utilizing of the Waste" (waste in this case referring specially to the small sizes of anthracite heretofore thrown away), my attention was naturally directed to the mash economical method of burning these small sizes.* Although I had been Familiar with the burning of anthracite for • many years, I found when I came to study the matter carefully that I was far from thoroughly understanding it, particularly with reference to the hurning of the smaller sizes, and that the literature of the subject, which was not very extensive, had more of a tendency to confuse than to enlighten me. I began therefore to investigate the subject so as to arrive if possible at some definite principles applicable to the question. With the aid of my assistant, Mr. John R. Wagner, I began at Drifton a series of experiments upon the burning of small anthracites. We first erected a small furnace with a grate 3 feet long and 2 feet wide, with which we experimented about two months. From the information thus acquired, we designed and built another furnace of about the same size, which gave us facilities for extending our investigations still further. After experimenting several months with the second furnace, we built a third one, which was much larger, having a grate area of about 68 square feet, which was exactly the size of the grate used in the boiler-plants of' Coxe Bros. &

Jan 1, 1894

By Adam J. Szladow

"INTRODUCTIONThe low calorific value and the high moisture and ash contents of run-of-mine (r.o.rn.) lignite limits the utilization of Saskatchewan lignites to thermal power plants in Saskatchewan, Manitoba and western Ontario. On the other hand, the low sulphur content and the low mining cost make lignites potentially attractive for new coal markets, if suitable processes can be found for upgrading. Economic upgrading to a higher quality fuel would make lignites more acceptable for industrial and utility boilers.This paper presents a summary of the process engineering evaluation of coal upgrading technologies conducted in support of a study of selected lignite upgrading processes. The summary presents the relative advantages and offers specific comments on the application of coal upgrading technologies for lignites. ResourcesLignite comprises the largest fossil fuel energy deposits in Saskatchewan. The total resources of immediate and future interest exceed 27 billion tonnes. The measured and indicated lignite resources of immediate interest are large enough to last for over 200 years at the current production rate of about 10 million tonnes per year (Table 1).Saskatchewan lignitedeposits are present in three different geologic formations: (I) the Ravenscrag Formation of Paleocene Age in southern Saskatchewan; (2) the Swan River Formation of Lower Cretaceous Age in central north Saskatchewan; and (3) the Judith River Formation of Upper Cretaceous Age in central west Saskatchewan. The deposits in the Swan River and Judith River Formation are not currently of economic interest.The Ravenscrag deposits are confmed to four lignite basins, namely: the Estevan, the Willow Bunch, the Wood Mountain and the Cypress basin (Fig. 1). The estimated resources of immediate interest in the four basins are over 7 billion tonnes (Table I). These resources include lignite seams at least 1.5 m thick, at depths of less than 45 m.The average quality of lignite varies relatively little between the four basins considering the distance between deposits (Table 2). There is, however, a progressive increase in moisture and ash content, and a decrease in calorific value and carbon content as one moves westward from the Estevan to Wood Mountain and Cypress basins. The Estevan area deposits are of higher rank despite the fact that they are stratigraphically younger than the Willow Bunch deposits (Cameron and Marconi, 1979).In upgrading lignite to a higher quality fuel, it is important to know not only the average proximate and ultimate analysis (Table 2) but also other lignite characteristics influencing the product quality . For example, Figure 2 shows the effect of lignite HIC ratio on the product equilibrium moisture for hydrothermal treatment of lignite. It is apparent that lignites with a higher HIC ratio, or high liptinite content yield much better quality product for the same hydrotreatment conditions. Other examples of lignite characteristics which are important in upgrading include distribution of sulphur forms and mineral matter, maceral composition and degree of surface oxidation."

Jan 1, 1989

By W. H. Ode, W. A. Selvig

"INTR0DUCTIONAccording to the Standard Specifications for Classification of Coals by Rank of the American Society for Testing Materials (A.S.T.M. Designation: D388-38).1/ high-volatile bituminous coals, subbituminous coals, and lignites are classified on the basis of calorific values on the ""moist"", mineral-matter-free basis. The moist basis refers to coal containing its natural bed moisture but not including visible water on the surface of the coal. It is assumed that the natural bed moisture of coal in the ground represents the moisture-holding capacity of the coal when in equilibrium with air at approximately 100 percent relative humidity. This moisture is sometimes referred to as inherent moisture as opposed to extraneous moisture, which wets the surface of the coal.To simplify procedures for determining the moisture-holding capacity of coal it is assumed, for all practical purposes that the moisture content of coal when brought to a standard condition of moisture equilibrium in an atmosphere of 97 percent relative humidity and 300 C. represents the bed moisture of the coal. The A.3.T.M. specifications provide that this condition of moisture equilibrium shall be obtained by bringing the coal to a standard condition of moisture equilibrium at 300 C. in a vacuum desiccator containing a saturated solution of potassium sulfate (97 percent relative humidity) according to the procedure proposed by Stanfield and Gilbart. 4/ No mention is made in the specifications for bringing partly air-dried samples to their natural bed moisture condition probably because of the hysteresis effect that has been noted by various investigators. 5/ In this connection it has been found from sorption studies that the adsorption and desorption curves for lignite are not entirely reversible. The hysteresis effect has been explained by some investigators as resulting from shrinkage of the lignite during drying, which causes a collapse of some of the capillaries, so the dried lignite cannot take up as much water as it held originally."

Apr 1, 1953

By Paul Torda

El presente texto analiza la posición de la institución educacional como "depósito" de talentos de ingeniería para futuras exploraciones dentro de posibles descubrimientos tecnológicos. En este sentido, la creación de un programa educacional adecuado genera el cultivo de una imagen del talento de ingeniería para los roles técnicos no tradicionales en la industria, como también para el gobierno local y el federal. El plan general involucra no solamente una ruptura con los procedimiientos tradicionales de clases, sino también origina nuevas demandas sobre los estudiantes por la independencia e iniciativa.

Nov 1, 1973

By R. K. Clifford

Chalcopyrite (CuFeSz) concentrates produced on the Lead Belt of Missouri typically contain 28 pct Cu and from 0.5 to 1.5 pct each of nickel and cobalt. Most of the nickel and cobalt are present as the mineral siegenite, (Ni,Co)3S4, almost none of which is recovered presently. The Bureau of Mines, U.S. Department of the Interior, has been conducting research on the chalcopyrite to produce a marketable fraction enriched in cobalt and nickel. Ferric chloride leaching of various chalcopyrite (copper) concentrates has yielded residues containing from 10.2 to 16.2 pct Ni and 8.3 to 11.4 pct Co with recoveries of approximately 79 pct. In another approach, laboratory froth-flotation tests have been conducted to investigate the effects of regrinding, roasting, baking, and steaming prior to flotation, with a variety of reagents. The best laboratory results were obtained after regrinding the copper concentrate to 94 pct minus 400 mesh and floating the chalcopyrite at natural pH (6.5) with 0.48 pound of sodium diethyldithiophosphate and 0.01 pound of sodium cyanide per ton of concentrate. This procedure yielded a tailings assaying 20.4 pct Cu, 5.2 pct Ni, and 3.45 pct Co. The tailings contained 67.2 pct of the nickel and cobalt and 15.9 pct of the copper originally in the copper concentrate; the remainder of the values reported to the flotation concentrate, which analyzed 29.6 pct Cu. To verify the laboratory flotation results, a 60 kg/hr process demonstration unit (PDU) was constructed in the Cominco American Magmont mill, Bixby, Mo.

Jan 1, 1978

By John E. Pahlman

The Bureau of Mines has conducted research on a dual leaching method that offers the potential for the economic recovery of the silver, as well as of the manganese, contained in some domestic manganiferous silver deposits. The method consists of an initial leach of manganese with aqueous sulfur dioxide (S02) followed first by a neutralization rinse, and then by a second leach with cyanide solution for the extraction of silver. Refractory silver in these ores was either (1) intimately associated with the manganese, (2) chemically resistant to cyanide leaching, or (3) rendered inaccessible to cyanide leach solution by the surrounding manganese mineralization. Less than 10 pct of the silver was extracted when minus 2.5-plus 1.3-cm ore pieces were leached directly with cyanide. However, extractions of 25 to 82 pct of the silver and 65 to 96 pet of the manganese were attained in laboratory column-leach tests using the dual leaching method on equivalent-sized ore pieces from seven domestic deposits.

Jan 1, 1987

By A. E. Raddatz

The Federal Bureau of Mines, as part of its goal to minimize undesirable environmental impacts associated with mineral processing, has investigated a soda-smelting technique for processing ilmenite into pig iron and a titanium slag that is suitable for a feed to a sulfate-process titania pigment plant. By this procedure, the generation of ferrous sulfate waste and its associated disposal problem is obviated. Molten titania slag was quenched, water leached, and filtered to yield a sodium titanate intermediate product containing about 70 wt-pct Ti02, 1 wt-pct Fe, and 20 wt-pct Na20. This intermediate product, high-soda titanate (HST), was mixed with H2S04, calcined, quenched, and water leached to produce a low-soda titanate (LST) product that contained only 2 wt-pct Na20. Both the HST and LST were evaluated as feed materials for preparing titania by the sulfate process. About 95 pct of the Ti02 was extracted from HST and LST by digestion in concentrated or 50-wt-pct-H2S204 solutions. About 95 pct of the solubilized Ti02 was subsequently recovered from the titanyl sulfate solutions by hydrolysis. Barren solutions from hydrolysis, containing about 26-wt-pct H2S04, were upgraded by vacuum-thermal evaporation or electrodialysis. The regenerated acid solutions were used to digest additional HST and LST.

Jan 1, 1979

By Edward C. Mabbs

U.S. industry must increase its productivity or lose its competitive position in the rapidly expanding and inter-dependent world markets. Recent improvements in material handling techniques and equipment, coordinated with production processes, provide one of the greatest opportunities to increase productivity. Successful managers must learn how to take advantage of these material handling opportunities to gain competitive position.

Jan 1, 1973

By Robert Lee

"International competition is presenting the Canadian steel industry with a major challenge. While Canadian companies have invested in new technology, research and development, activity by Canadian industry, in general, lags significantly behind that of its major competitors. While OECD countries increased their research and development spending during the 1980s, Canada's R&D funding levels have changed little since the early 1970s and have actually dropped since 1986, despite greater public discussion on the issue. This fact is reflected in the relatively slower pace of innovation in the Canadian steel industry, despite the efforts of both government and industry.In addition, there are a number of factors that undermine the competitiveness of the Canadian steel industry, in particular in relation to the United States. Among these are higher interest rates, labour costs and tax rates, as well as an overvalued Canadian dollar. Canadian steel companies are thus feeling the full impact of the current economic recession. This recession, combined with recent strikes, have, for example, caused serious losses for Algoma and Stelco.In spite of these objective difficulties, the Canadian steel industry has been able to survive and continues to strive for a leadership position in the world market. Since World War II, heavy investment in new technology has allowed the industry to upgrade its facilities and increase productivity, while at the same time improving quality and reducing pollution. This last point is particularly important in view of public insistence on improvements to the quality of the environment. Over the next decade, this technological innovation must be maintained and intensified if industry is to meet the challenge of international competition and attain a leadership position."

Jan 1, 1991

By Jaime V. Ongpin

"The author will give a capsule history of Benguet Corporation and attempt to identify its operating environment and underlying ""strategy"" until 1974, when the company's management assumed office. The company's operating environment and ""interim strategy"" from 1975 to 1980 will be described. The discussion will cover an environmental forecast and define the company's"" strategy for survival"" for the balance of this century.IntroductionIn December last year, I was asked to deliver a pa per for the management lecture series a t the Asian Institute of Management in Manila. The paper was entitled "" The 1970's: A Decade of Disaster For New Mining Projects"" . In that paper, it was my task to review the results of new mining projects in the Philippines during the previous decade. The task was not a happy one, for the results themselves were not only unhappy, but disastrous on a scale unprecedented in Philippine history. Out of a total of 17 major mining project s involving an aggregate capital investment in excess o f US $2 billion during the 1970s, only three project s have been profitable, and all the rest have either gone "" belly up"" or are teetering on the very edge of financial disaster.It is in the context of this shocking record of widespread financial dislocation in the Philippine mining industry, (which in varying degrees appears to have been paralleled in virtually every other country in the world where mining is an important sector of the economy) that I address myself tonight to the theme of this symposium. My paper is entitled ""Risk Management Strategy in Mining"" and it is sub titled ""Can Benguet Corporation Survive to Celebrate its Centennial in 2003?"". I have inserted the word ""Strategy"" and added the subtitle, because I wish to make it clear first, that my presentation to you tonight will be focused exclusively on the strategic aspect s of risk management, and second that I will limit myself to Benguet's own ""in-house"" experience and perspectives on this subject. I would caution you therefore, to consider my views within this restricted framework, and in light of the particular and sometimes peculiar circumstances within which Benguet Corporation operates."

Jan 1, 1983

By Otto Wartenweiler

AFTER the phenomenal development of the new mine, the United Eastern Mining Co., with Mr. Frank A. Keith as President, decided to install a reduction plant. The character of the ore, closely resembling the product of other mines in the district, did not offer any particular metallurgical problems, since these neighboring mines had been operating milling plants very successfully for several years. Thus, with the general metallurgical treatment fairly well-defined, the matter of next importance was to design and construct a plant as economically as possible, that would give the best results obtainable. The ore consists of a mixture of calcite and quartz in about equal quantities with some decomposed andesite, while the average value of mill heads were calculated to be $20 plus in gold, no silver being present. The tonnage to be treated at ,the start was fixed at 200 tons, but, in view of the favorable condition of the mine, it was considered advisable to arrange facilities, for increasing the capacity to 400 tons. . While it is best in some cases to adopt a distinct independent unit system where future enlargement is contemplated, in this instance the plans worked out differently. Summarized, the treatment adopted consists of coarse crushing, coarse grinding, fine grinding and cyanidation. For coarse crushing the selection of the No. 6 size gyratory crusher was governed not so much by its capacity as by the size of the feed openings to properly receive the run-of-mine material, which was too coarse for a No. 5 crusher. To accomplish the coarse grinding to. best advantage, a careful comparison was made between the modern steel ball mill-the last word in crushing and grinding-and one of its older, and, at that time better-established, competitors, the rolls.' The following conclusion was reached: While it is possible that for the sized plant under consideration the cost of operating rolls might be a little cheaper, it is preferable to use ball mills because a simpler arrangement is possible, and1 because. the handling of, ball mills is preferred by the operator.

Jan 11, 1917

By Robert H. Bedford

As a result of the Bureau's general survey of quicksilver deposits in California, the Guadalupe mine area was among those chosen for exploration. Exploration, such as that at the Guadalupe, was aimed primarily at maintaining or increasing production of quicksilver during the critical wartime period. The project as originally planned contemplated sampling the dumps and prospecting some promising surface exposures by bulldozer trenches. Salvage was the objective in the first case and new ore in the second. Later, a program. of diamond drilling was added, designed to add both to ore reserves and to geological information. This report summarizes the history, notes the location and physical features, mentions property and ownership, and presents a brief description of the mine. The geology is outlined, and the work done by the Bureau of Mines is explained and supplemented with accompanying maps. The work involved repairing 1-1/4 miles of roads; bulldozing 13 trenches by removing 2,345 cubic yards of dumps to provide for sampling; digging and sampling 9,284 linear feet of hand trenches, digging and sampling numerous test pits; and drilling 7 diamond-drill holes, totaling 1,000 feet. A total of 419 samples, including 10 percent check samples, were taken. Exploration work in California at the time of this project was under the direction of Spangler Ricker, supervising engineer of the California field office, the late Robert Bedford was in charge of the fieldwork on this project.

Jan 1, 1950