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By Dean H. Ambrose

Computer programs are presented that provide mine electrical systems based on computer modeling for design and safety analysis suitable for large or small computer systems and handheld calculators. Bureau of Mines research has resulted in load-flow, fault, grounding, reliability, short-circuit, transient, and cable ampacity computer models that enable mine power system engineers to analyze their system or to design a system. This report describes the capabilities of the program.

Jan 1, 1984

"The technical report form provides the headings to be used and content required when a completing a technical report to be filed in accordance with NI 43-101. The technical report is meant to be a summary of the scientific and technical information on the mineral property. The Instrument (NI 43-101) requires that technical reports be filed with Canadian provincial or territorial securities regulators to support certain disclosure of mineral exploration, development and production activities and results in order to permit the public and analysts to have access to information that will assist them in making investment decisions and recommendations. FORM 43-101F1 TECHNICAL REPORT INSTRUCTIONS (1) The objective of the technical report is to provide a summary of scientific and technical information concerning mineral exploration, development and production activities on a mineral property that is material to an issuer.This Form sets out specific requirements for the preparation and contents of a technical report. (2) Terms used in this Form that are defined or interpreted in National Instrument 43-101 Standards of Disclosure for Mineral Projects (the “Instrument”) will bear that definition or interpretation. In addition, a general definitioninstrument has been adopted as National Instrument 14-101 Definitions that contains definitions of certain terms used in more than one national instrument. Readers of this Form should review both these national instruments for defined terms.(3) The qualified person preparing the technical report must use all of the headings of the items in this Formandmay create sub-headings. If unique or infrequently used technical terms are required, clear and concise explanations must be included. (4) No disclosure need be given in respect of inapplicable items and, unless otherwise required by this Form,negative answers to itemsmay be omitted. Disclosure included under one heading is not required to berepeated under another heading. (5) The technical report is not required to include the information required in Items 6 through 11 of this Form tothe extent that the required information has been previously filed in a technical report for the property being reported on, the previous technical report is referred to in the technical report and there has not been any material change in the information. (6) The technical report for development properties and production properties may summarize the information required in the items of this Form, except for Item 25, provided that the summary includes the materialinformation necessary to understand the project at its current stage of development or production. (7) The technical report may only contain disclaimers that are in accordancewith section 6.4 of the Instrument and Item 5 of this Form."

Jan 1, 2003

By P. R. Smith, R. A. Walters

Borates have been mined in the desert areas of California and Nevada for more than 100 years. To about 1890, playa surface mining provided the chief sources of boron minerals. Underground mining of colemanite and later of borax and kernite was predominate until about twenty years ago. Open pit mining of the large deposits of borax and kernite near Boron, California has been most significant for the past twenty years. Mining of colemanite in the Ryan, California area, near Death Valley, began in 1907. Following the discovery of the large deposits in the Boron area (about 1957), mining in the Death Valley area became nearly nonexistent. Only small tonnages were mined for special uses. Little mining was done in the Boraxo area near Ryan. The first claim was made in about 1915. In 1960 the area became the property of the Kern County Land Company, which was acquired by Tenneco Oil Company in 1967. In 1976 the various boiate properties and claims in this region were acquired by American Borate Corporation. The open pit mine is now approximately 122 m (400 ft) deep, 910 m (3000 ft) long and 305 m (1000 ft wide). The borates in the Boraxo pit consist primarily of three minerals. These are about 50% colemanite (CB6011 5H20), about 40% probertite (NaCaB50g 5H20), and 10% ulexite (NaCaBgOg 5H20). The colemanite, along with boric acid and high-grade colemanite ore from Turkey provide the only sodium-free borates for production of textile grade fiberglas. When heated to its decomposition temperature, colemanite decrepitates to a fine powder, which is the basis for the concentration process. The gangue minerals in this deposit are primarily calcite and clays, including bentonite. The ore body has a very low arsenic content, which is a desirable feature. Test work had been done with samples prior to the results discussed herein. This paper will discuss results of test work which were the basis for erection of a plant, and the subsequent plant operations. Laboratory Calcination and Air Tabling Tests Laboratory calcination tests showed that substantial upgrading of the borate could be accomplished by calcining followed by screening of the calcined material. Removal of the + 28 mesh calcine resulted in borate losses of less than 10% with a rejection of 40 weight % or more of the calcine. The minus 65 mesh calcine generally met the requirement of containing 48%, or more, B203. The minus 28 plus 65 mesh material contained an intermediate quantity of borate and would require additional treatment. Testing demonstrated that ore would not have to be reduced to a size finer than 19 mm (3/4 in.) prior to calcination. A temperature range from 400 to 455OC (750 to 850°F) was apparently satisfactory. Calcination at a temperature of 48Z°C (900°F), or higher, was unsatisfactory due to fusion. All laboratory calcination tests were static tests conducted by placing small covered charges in a laboratory furnace for 40 min. In all tests vapor issues from the furnace for 5 to 7 min. Following this period the ore could be heard "popping," due to decrepitation of the colemanite. The reaction generally continued for approximately one-half hour. Various size fractions of the calcination products from laboratory tests were subjected to laboratory air tabling tests, usually after removing the plus 28 mesh material. Laboratory air tabling tests were conducted employing a Whippet V-80 model air table manufactured by Sutton, Steele and Steele Co. now known as Tripple S Dynamics. Variables include both end and side-tilt, speed of vibration, and quantity of air rising through the deck. In addition to the variables in the machine itself, the feed rate is also a rather critical variable. Testing demonstrated that all - 28 mesh size fractions of the calcine could be successfully concentrated to 48% F2O3 or greater. For the finer material recoveries into the concentrate were between 85 and 90% of the borate. With the coarser material a substantial amount of middling was produced which required cleaner tabling. Laboratory calcination and air tabling tests indicated a process whereby the borate could be concentrated to about 50% B203 with borate recoveries approaching 90%. Moreover, the iron content of the concentrate was well below the required specification of 0.3% Fe2O3. Pilot Plant Calcination Following the laboratory test work described above, pilot plant testing was conducted to prove the process, provide data for engineering studies, and provide product for a prospective purchaser. The kiln used was 0.9 m-diam (3 ft) by 9.0 m (30 ft) long and had a belly section 1.2 m-diam (4 ft) by 2.74 m (9 ft) long near the discharge end. The kiln was operated at a speed of 0.7 rpm. Gas was fired into the kiln at an average rate of 27.1 m3/hr (958.4 cu ft per hr). The air to gas ratio used was 10:1. The ore was fed to the kiln countercurrent to the flame and discharged through a hopper into a screw conveyor which discharged to a 1.2 m (48 in.) Sweco separator. The separator had 28, 65, and 150 mesh screen cloths, with the plus 28 mesh fraction being discarded. The minus 28 mesh fractions were later subjected to air tabling. The exit gases, containing some calcine dust, were swept through two cyclones to recover the dust. The gases then were scrubbed in a Ducon scrubber; very little dust reported past the first cyclone. The dust from the first cyclone was also saved in drums. In addition to the gas rate, the flue gas velocity, after

Jan 1, 1981

By M Campbell, A Jeffcoate, J-C Vannay, B Perriaux

Phyllite waste rock at the Simandou Pic de Fon and Oueleba Iron Ore deposits (Guinea, West Africa) variably weather to form clay-rich rocks. This weathering modifies the mineralogical, geochemical and mechanical properties of the phyllites. Final pit walls will largely be formed in phyllite, so it is critical that the slope design takes account of the distribution and variability of weathering within this material. Phyllite intact material strength has been correlated with geochemistry (ie mineralogy) facilitating the establishment of domains of intact rock strength (IRS). In order to calibrate this geochemically-based strength model, core from diamond drilling (DD), comprising a variety of phyllite material types (from soft clay to moderate strength rock), was geotechnically logged to assess intact strength; further validation was achieved by uniaxial compression (UCS) testing and point load tests (PLT). The correlation between geochemistry and strength offers a potentially powerful tool for differentiating basic mechanical properties of the rock mass, especially at Simandou where a significant proportion of the resource drilling is done by reverse circulation (RC). Unlike DD, reliable measurements of material strength from RC samples can prove difficult, especially in clay-rich material. Therefore, gaining information from routine geochemical assay results, independent of the sampleÆs physical integrity, is a valuable and cost-effective source of reliable empirical data. Data acquired using this approach provides an opportunity to develop strength domains, which are integrated into mine planning block models. In combination with other geotechnical characteristics, this can be a valuable input to the development of geotechnical models. This technique has allowed spatially representative rock strength data to be derived and offers the potential for optimising pit slope angles in areas where hitherto conservative design criteria had been assumed.CITATION:Jeffcoate, A, Campbell, M, Vannay, J-C and Perriaux, B, 2013. Modelling intact rock strength of clay-rich rocks with geochemistry at the Simandou iron ore deposit, Guinea, West Africa, in Proceedings Iron Ore 2013 , pp 193-202 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

By T. Grammatikopoulos, C. Gibson, M. Aghamirian

"Lithium is a silver-white alkali metal that has recently become a target commodity for many junior mining companies. Lithium has a wide range of applications including ceramics and glass, lubricating greases, air treatment, continuous cast mold flux powders, polymer production, and primary aluminium production, but the current popularity of lithium can be attributed to an anticipated increase in demand for lithium ion batteries. The most common lithium-bearing minerals occur globally in pegmatites and granites. Spodumene [LiAl(Si2O6)] is the most desirable economic lithium-bearing mineral due to its high lithium content (theoretical Li2O content of about 8 wt%). However, lithium can also be hosted in petalite, micas, tourmaline, and phosphate minerals. In Canada there are a number of spodumene deposits in Ontario and Quebec at development stage that have the potential to help narrow the forecast lithium supply gap stemming from the widespread acceptance and use of electric vehicles. This paper discusses the beneficiation of hard rock lithium deposits, including an in depth review of the use of flotation to upgrade spodumene. A review of the use of dense media separation (DMS), magnetic separation, and roasting as means to process spodumene bearing ores is also discussed. INTRODUCTION Lithium is a soft, silver-white alkali metal with atomic number 3. As of 2015, the end use of lithium was estimated to be as follows: lithium-ion batteries (35%), ceramics and glass (32%), lubricating greases (9%), air treatment (5%), continuous casting mould flux powders (5%), polymer production (4%), primary aluminium production (1%) and other uses (9%) (Jaskula, 2014). Lithium is found in brine lake deposits containing lithium chloride and in hard rock pegmatite deposits. It can be sold in a range of product types; as mineral concentrate, brine, compound (lithium carbonate or lithium hydroxide) and metal depending on its intended end use (Goonan, 2012; Jaskula 2014). In the 1990s the United States and Chile were the largest producers of lithium globally, primarily from brine operations (Ober, 1994). Most of the lithium compounds produced over this period were consumed in the glass, ceramics and aluminium industries. In the 1980s and 1990s there were two major lithium operations in the United States: brine operations at Silver Peak, Nevada, and a spodumene mine and lithium carbonate/lithium hydroxide plant in Kings Mountain, North Carolina. The Kings Mountain operations ceased production in 1991 and were eventually dismantled in 1994. Brine operations were also active in Argentina and Chile, while the largest lithium ore deposit in the world was operated at Greenbushes, Western Australia. During this period it was predicted that major shifts in demand for lithium would come from the success and widespread use of electric vehicles (Ober, 1994)."

Jan 1, 2017

By William Crowe Kellogg

By now the term AFMAG has been fairly well introduced to the mining industry. Several papers have been presented and published concerning AFMAG (references 1 and 2). However, the term AFMAG remains rather new to the industry, and the exact place which AFMAG occupies in the minerals exploration field is still somewhat uncertain in the minds of many who may have use for this new and interesting method of electromagnetic exploration for conductive minerals. It is the purpose of this paper to again review the basic theory of the AFMAG system, to describe and - illustrate the equipment and data presentation as used in aerial surveying, and to provide some idea of the cost and nature of airborne AFMAG surveys. This paper is primarily of a general descriptive nature and presents AFMAG theory in a schematic rather than mathematical form; emphasis is placed on the equipment, the present method of data presentation and survey operations, since it is believed that such, information will now generally best serve to inform the mining industry of the characteristics of the AFMAG system.

Jan 1, 1960

By Charles L. Allie

In June, 1980, Reserve Mining Company completed a 370 million dollar construction project which included process modifications required to reduce the silica content from nine to five percent by weight and increase the iron con- tent of Reserve's pellets from 62.5 to 65.8 percent by weight (dry basis); process modifications and additions required to implement "best available technology" water quality controls; and installation of "best available technology" air emission controls. The new process began operation on June 24, 1980 and is being optimized to meet design objectives. Dry cobbing, fine screening and flotation additions to the concentrating process have reduced the silica content of Reserve's pellet from nine percent to five percent by weight. The process reclaims and reuses all process waters, stores all tailings in a zero discharge tailings basin and meets all air emission standards.

Jan 1, 1980

By M. Yekich, W. R. Reed, G. Ross, A. Salem

"Canopy air curtain (CAC) technology has been developed by the National Institute for Occupational Safety and Health (NIOSH) for use on continuous miners and subsequently roof bolting machines in underground coal mines to protect operators of these machines from overexposure to respirator coal mine dust. The next logical progression is to develop a CAC for shuttle cars to protect operators from the same overexposures. NIOSH awarded a contract to Marshall University and J.H. Fletcher to develop the shuttle car CAC. NIOSH conducted laboratory testing to determine the dust control efficiency of the shuttle car CAC. Testing was conducted on two different cab configurations: a center drive similar to that on a Joy 10SC32AA cab model and an end drive similar to that on a Joy 10SC32AB cab model. Three different ventilation velocities were tested—0.61, 2.0, 4.3 m/s (120, 400, and 850 fpm). The lowest, 0.61 m/s (120 fpm), represented the ventilation velocity encountered during loading by the continuous miner, while the 4.3 m/s (850 fpm) velocity represented ventilation velocity airflow over the shuttle car while tramming against ventilation airflow. Test results showed an average of the dust control efficiencies ranging from 74–83% for 0.61 m/s (120 fpm), 39–43% for 2.0 m/s (400 fpm), and 6–16% for 4.3 m/s (850 fpm). Minor modifications to the shuttle car CAC design and placement improved the dust control efficiency to 51–55% for 4.3 m/s (850 fpm) with minimal impact on dust control efficiencies for lower ventilation velocities. These laboratory tests demonstrate that the newly developed shuttle car CAC has the potential to successfully protect shuttle car operators from coal mine respirable dust overexposures. INTRODUCTION The development of the canopy air curtain (CAC) dates back to the 1970s starting with the initial development of the CAC by the Donaldson Company, Inc. under contract from the U.S. Bureau of Mines [Krisko 1975]. This CAC was originally developed for continuous miner operators when continuous mining machines had cabs. The need for a CAC on the continuous miner was eliminated when the cab was removed from the machine design. However, CAC development progressed to include CAC designs for a roof bolting machine to protect roof bolters from respirable coal mine dust [Goodman and Organiscak 2002, Listak and Beck 2012, Reed et al. 2017]. This roof bolting machine CAC research continues to the present day."

Jan 1, 2018

Studies were carried out to examine the effect of introducing varying levels of a pisolitic limonite (Yandi) into an Australian iron ore blend. Compared with other ores in the blend Yandi is more porous. This structure is found to have a considerable effect on the sintering behaviour of the blends. Pilot-scale sintering studies and plant trials showed that sinter plant productivity, sinter reducibility and the low-temperature reduction degradation index altered significantly when the level of Yandi in the ore blend exceeded 15 wt%. The improvement in the reducibility is thought to be a direct result of increased porosity in the sinter and the presence of highly reducible pseudo-relict haematite. Plant trials have also shown that operational changes could be made to the sinter plant to avoid potential deterioration of productivity

Dec 1, 1991

By John G. Tate, George W. Nicolson, James L. Bruce

For more than 25 years modern mining has been carried on in the Island of Cyprus, Mediterranean Sea, by the Cyprus Mines Corp. of Los Angeles, Calif. The general features of these operations have been described.' From 1935 to June 30, 1940, top slicing in the Mavrovouni mine produced in excess of 2,000,000 long tons. The rate for the last part of this period was approximately 50,000 tons per month. Description of Ore Bodies The shape Of the Ore body now being top sliced in the Mavrovouni mine is an inclined irregular pyritic pipe or chimney of large dimensions) extending from unknown depths below the bottom level at elevation 250 ft below sea level) to the highest level at elevation 400 ft above sea level. The present surface above the ore body lies between 550 and 650 ft elevation. Undoubtedly the Ore body extended to surface and was Mined by the ancients by opencast and underground methods. The old pits were gradually filled up by erosion until now they show as slight depressions in the topography. The largest area of the Ore body is at elevation 170 ft below sea level, where the horizontal section is almost elliptical, with a width of 600 ft and a length of 1000 ft. Below this elevation the shape of the ore body continues to be elliptical but with reduced dimensions. Above this elevation it rises flatly towards the south with a substantial depression in the upper-side of the chimney and irregular ridges or wings rising on each side. It gradually grows smaller as it rises. The ore body is a pyritic replacement of andesitic extrusives which occur as flows and pillow lavas. With the exception of the upper portions of one or two small offshoots rising from the main ore body, all ore is completely surrounded by andesite. Most of the andesite lying above the ore body is covered by marl and Clay sedimentaries of Miocene and later age, The Miocene is presumed to be older than the ore mineralization. That portion of the andesites which lies between the sedimentaries and the ore body has been broken by slumpage, and weathered, and iron-Stained by the descending meteoric waters, It is a mass of crushed material that caves and follows downward readily as the top-slice stopes retreat. The main ore body is a heavy cupreous pyrites mass assaying on the average about 4.0 pet Cu; 48.0, S; and 42.0, Fe. In places the copper drops below 1.0 pet. on the borders of this heavy sulphide ore body lie disseminated cupreous pyrites ore deposits of large dimensions. some portions of these appear to be closely related to the heavy sulphide ore, as the ratio of copper to pyrite is about the same. The greater part of the disseminated ore, most of which lies immediately to the east of the main Ore body, contains less than 0.10 pct Cu and about 25.0 to 30.0 pct S. 'This probably can

Jan 1, 1949

By J. P. Norrie

The presence of copper-bearing rocks in the Coppermine River area of Arctic Canada has long been known. Word of this was probably circulated along the Arctic and Hudson Bay coasts by the Eskimo, who no doubt traded articles made from native copper among themselves. It is not known whether any of these ever reached the southern shores of Hudson bay. The Indians, also, knew of the occurrence of copper here, as they served as guides to Hearne when he made a successful trip over land from Hudson bay to the Coppermine river. The possible commercial importance of these copper deposits was early appreciated by the Hudson's Bay Company, who on December 7th, 1770, despatched Samuel Hearne from Fort Prince of Wales, now Fort Churchill, to investigate them. He reached the then unnamed river at a point forty miles from its mouth on July 14th, 1771, and after surveying the river to its mouth and naming it the Coppermine, he retraced his steps to Fort Churchill by an overland route. Hearne's expedition was probably the most remarkable and arduous ever attempted in Arctic exploration, but the actual value derived from the trip was entirely out of proportion with the hardships he must have suffered. He verified the occurrence of native copper, but his astronomical and other observations were of little value. Until 1912 no concerted effort was made to obtain any definite information as to the geology and economic value of the Copper-mine deposits.

Jan 1, 1931

By E. I. Gedgagov, A. V. Tarasov, V. M. Paretsky, A. D. Besser

"Hydrometallurgical technology and equipment have been developed in the Gintsvetmet Institute (Moscow, Russia) for processing of rhenium-containing alloys waste to produce commercial- grade metal products with recovery of all valuable components. The technology is based on waste treatment with hydrofluoric acid with subsequent decomposition with nitric acid at pH = 1 to 1.5 to ensure maximum degree of rhenium conversion to the liquid phase of pulp. Some experimental lots of ammonium rhenate and ammonium molybdate were produced at a pilot plant. The following main techno-economic performance values have been achieved when processing grinding residues (balance-based tests were conducted with a lot of SP-2 alloy waste) to produce ammonium rhenate of high purity, as well as tantalum-niobium, nickel-cobalt and tungsten-molybdenum concentrates: rhenium recovery of 92.3% into a commercial product from grinding waste; production cost of 170-210 US Dollars per 1 kg of rhenium in commercial-grade ammonium perrhenate taking into account its recovery; tantalum and niobium recovery of 75- 77% into bulk concentrate; nickel and cobalt recovery of 88-89% into a mixed concentrate; tungsten and molybdenum recovery of 88% to 91% into concentrate. Introduction During the recent years the use of rhenium has been increasing at a fast pace worldwide [1]. The growing prices for this refractory metal are an essential factor encouraging collection and processing of rhenium-containing wastes. Double-component waste, such as molybdenum- rhenium or tungsten-rhenium alloys can be readily decomposed and the process of rhenium recovery does not cause, therefore, any difficulties. The main separation of metals takes place at the stage of waste digestion with acid, because molybdenum and especially tungsten remain in precipitate in the form of hydrated oxides and rhenium reports in the liquid phase, from which it can be easily extracted by solid ion-exchange or solvent extraction techniques. However, most of rhenium is contained in super-alloy waste [2]. Such waste has different contents of valuable components, such as tantalum, niobium, molybdenum, tungsten, cobalt, nickel, chromium and other metals. The basis of all such super-alloy waste is nickel. Refractory metals in the super- alloy composition are in the form of metallized target-oriented eutectics, and due to this reason hydrometallurgical decomposition of such alloys is a difficult technological task."

Jan 1, 2008

By P. M. Solozhenkin, O. I. Solozhenkin

Currently lead nitrate is used as the recommended activator in Au-Sb flotation. However, due to high levels of hazard, the application of lead nitrate constitutes a potential risk for human health and ecosystems. To mitigate the harmful impact of activators on the environment, further research on activation by alternative reagents has been launched to promote the efficiency of antimony ores. In the present study, adsorption of lead, copper and zinc cations has been investigated by nuclear quadrupole resonance (NQR), which has revealed copper and zinc cations penetration into the mineral lattice, although lead cations remain adsorbed on the mineral surface. However, the combination of zinc and copper cations leads to preferential volumetric adsorption of zinc in the mineral lattice and further surface adsorption of copper which positively affects the activation process. It has been established that activation of stibnite can be achieved effectively with the mixture of Cu2+:Zn2+=1:1 as much as with lead cations in antimony ore flotation. As a part of this work, the computational modelling has been used in prognosis of reagent activity evaluation (PRAE) for stibnite and aurostibite flotation by sulfhydryl collectors. The lower PRAE value the stronger the ability of a collector to form a stable complex with a mineral. PRAE has shown a strong negative trend to complex formation on the mineral surface of aurostibite and stibnite. Based on these studies, several collectors have been proposed for effective flotation. To verify the feasibility of proposed collectors tetramethyl (ethyl) thiuram disulfides (TMTD and TETD) have been used for flotation of stibnite, sphalerite, pyrite and arsenopyrite. Flotation performance has been improved significantly with TMTD that allows the avoidance of lead nitrate application and increases the recovery of antimony by 8,16%. Keywords: Stibnite, aurostibite, activation, lead, copper, flotation reagents

Jan 1, 2020

By Gust Van Lysebetten, Monika de Vos

The objectives of monitoring geotechnical structures can be very diverse, including optimisation of the design, quality control, long term performance assessment, and approval of innovative execution techniques. In this context, new measurement techniques, based on optical fibres, offer the possibility of accurate strain and temperature monitoring with very small sensors and limited numbers of cables. This paper briefly presents optical fibre sensing techniques and illustrates their application to geotechnical structures by means of some projects, carried out by the Belgian Building Research Institute (BBRI).

May 1, 2022

By H. E. Powell

This Bureau of Mines study describes two processes for recovering zinc, copper, silver, and iron from the magnetic fraction of a waste generated by the primary smelting of zinc in horizontal retort distillation furnaces. The laboratory-scale procedures include reduction melting to produce a high-copper iron and matte smelting to produce a copper-silver-iron matte and metallic iron. Zinc is fumed in each process and recovered as the oxide. Direct reduction melting yields a metallic iron high in copper and a small rich matte from the sulfur contained in the waste. The products of the second process are a matte containing essentially all of the silver, about 80 pct of the copper and 37 pet of the iron, and an iron-copper alloy containing all of the remaining copper. Extraction with metallic lead or sodium sulfate recovers essentially all of the copper from the molten high-copper iron and yields an acceptable foundry iron.

Jan 1, 1973

By Adrian C. Dorenfeld, Theodore Balberyszski, Strathmore R. B. Cooke

This paper reports results of studies of sulfidiza-tion of base-metal oxides and silicates with gaseous sulfur, hydrogen sulfide gas and pyrite and of their subsequent flotation with xanthate collectors. It has been established that a 100% conversion of the oxides of lead, zinc, copper and manganese into their respective sulfides is possible and is a function of reaction time and temperature. Sulfidization of chrysocolla (with malachite and azurite inclusions) resulted in a product containing covellite, chalcocite and digenite in proportions depending on conditions of sulfidization. Reacting cassiterite with gaseous sulfur, H2S or pyrite at various reaction temperatures resulted in its partial conversion to a tin sulfide. Flotation of the products of sulfidization indicated that their behavior is similar to that of natural sulfide minerals, the recovery being a function of the con-version of the oxides to their respective sulfides. The gradual depletion of high-grade sulfide mineral deposits has turned the attention of the mineral industry to the recovery of metals from the oxides and silicates. Anionic (fatty acid) and cationic collectors have in some cases made flotation of nonsulfides possible, although, in general, flotation of oxide ores is not as yet an economic process. Consequently, many attempts have been made to sulfidize the surface of the relevant oxide mineral and thus make it amenable to to In recent years, Bautista and sollenberger4 have investigated the conversion of a number of oxides to sulfides by reacting them with molten sulfur. This paper presents the results of .investigations carried out at the School of Mineral and Metallurgical Engineering at the University of Minnesota to study the flotation characteristics of some sulfidized minerals. It is divided into two parts: The first describes the sulfidization of a number of oxides and a silicate, and the second presents the flotation behavior of the products of sulfidization. SULFIDIZATION OF METAL OXIDES AND SILICATES Experimental Procedure: SULFIDIZATION WITH HYDROGEN SULFIDE GAS — Sulfidization with H2S gas was carried out in a horizontal tube furnace (Fig. 1). The temperature of the furnace was regulated by means of a Variac and determined with a chromel-alumel thermocouple. The constant temperature zone in the furnace was about 6 in. in length, and it was within this zone that the experiments were carried out. The gases employed were supplied from standard compressed gas cylinders, and passed through a glass-wool filtering tower before entering the Pyrex reaction tube. Gaseous products were bubbled through a water tower (used as a flow indicator) and excess H2S was burned off at the end of the outlet tube. Samples of reagents or minerals were placed in a 3-in. Alundum boat and inserted within the constant temperature zone of the furnace. Nitrogen gas (6 ppm O2 content) was passed through the tubes at room temperature to remove air in the tube, and continued to pass while the furnace was heating up. When the required temperature was reached, H2S was allowed to enter the tube and the supply of nitrogen was shut off. At the end of the reaction period, the reverse procedure was carried out, the samples being cooled to room temperature in a nitrogen atmosphere. The flow rate of H2S gas was maintained fairly constant in all experiments, although the rate was not determined. SULFIDIZATION WITH SULFUR POWDER - To study the sulfidization of minerals with sulfur (liquid or gaseous), samples of the mineral, together with a measured quantity of sulfur powder, were placed in a stainless steel (SS316) reaction vessel (Fig. 2). The reaction vessel was heated in a vertical resistance coil furnace, the temperature being regulated through a Variac. At the end of the experiment the vessel was cooled to room temperature before the sample was removed. Cooled products were leached with CS2 to remove any free sulfur. SULFIDIZATION OF TIN CONCENTRATES - Fifty gram samples of sulfidized tin ore were prepared by mixing -48 mesh Bolivian tin concentrate with -48 mesh Ottawa sand in various proportions and sulfi-dizing these samples with H2S in the horizontal tube

Jan 1, 1969

By V. V. Pankov, W. R. Chen, L. R. Zhao

Diffusion nickel aluminide and overlay NiCrAlY coatings were deposited on single crystal Ni-base superalloy CMSX-4 substrate using the chemical vapour deposition and cathodic arc evaporation techniques, respectively. The NiCrAlY coating underwent a post-deposition heat treatment in an argon atmosphere at 1100°C for 2 hours. All the coated specimens were subjected to furnace cyclic oxidation testing at 1000°C up to 1500 one-hour cycles. The morphology, elemental composition and phase constituents of the oxide scales formed on the coating surface were analyzed using scanning electron microscopy, energy-dispersive spectroscopy and X-ray diffraction. It was observed that the diffusion aluminide coating experienced accelerated weight gain at an early stage of the test (150-400 cycles), whereas the overlay NiCrAlY coating fabricated by cathodic arc evaporation maintained steady-state weight gain throughout the entire 1500 cycle test duration. This difference in oxidation behaviour between the two types of coatings is attributed to the early development of nickel-aluminum spinel on the surface of the diffusion aluminide coating, in contrast to the stable aluminum oxide scale formed on the overlay NiCrAlY coating during cyclic oxidation testing.

Jan 1, 2015

By Peter A. Anderson

Innovative solution management is being used at Coeur-Rochester Inc. to increase bullion production and decrease overall costs. New operating programs call for reduced solution application rates to the heap that increase pregnant stream values at a lower and less costly plant flow rate. Silver and gold extractions are equal regardless of leach solution application rate. Solar heating is used in the summer months to increase the core temperature of the heap for use in the winter months to enhance the silver extraction rate. The Coeur-Rochester silver and gold mine, located in central Nevada, is an open pit, heap leach operation that has been in production since late summer of 1986. The facility produces about 15.4 kt/d (17,000 stpd) of -13 mm (-0.5 in.) ore to a leach area consisting of more than 37 hm2 (4 million sq ft) of 80 ml HDPE liner. Crushing is accomplished with a 1220x 1520-mm (48- x 60-in.) jaw crusher. This is followed by a single 2.1-m (7-ft) standard cone crusher and is completed with four 2.1 m (7 ft) short-head cone crushers. Agglomeration is not required for the crushed ore, which contains less than 10% -75. µm (-200 mesh) material. The ore contains 53 g/t (1. 55 oz per st) silver and 0.377 g/t (0.011 oz per st) gold. It is end dumped over previously leached material, leveled, and then ripped with a bulldozer. Drip emitters at 810 mm (32 in.) spacings on 910 mm (36 in.) centers are used to apply barren solution to the heap. A portion of the emitter lines are buried I m (3 ft) deep during the winter months and are placed on the surface the remainder of the year. The leach solution contains 0.75 g/L (750 ppm) sodium cyanide solution at a pH of 10.5 - 11. Dissolved oxygen averages 8 ppm in the pregnant solution. Cyanide consumption for the ore to date is just under 0.2 kg/t (0.4 lb per st). The pH control is accomplished by adding I kg/t (2.0 lb per st) 0-10 mm (-0.4 in.) pebble lime to the ore at the tertiary crusher. The pregnant solution is processed by a conventional closed circuit Merrill-Crowe system. The pregnant solution is clarified through three 158 L/sec (2500 gpm) US Filter Auto-jet clarifiers. This is followed by deareation through a

Jan 1, 1989

By Joan T. May

Federal Bureau of Mines studies showed that some organic amine hydrochlorides can be used as dehydrating agents for producing anhydrous magnesium chloride from hydrates and brines. Laboratory-scale tests showed that double salts are formed by reacting 35 percent magnesium chloride brine or spray-dried magnesium chloride with pyridine hydrochloride. Heating decomposed these double salts to anhydrous magnesium chloride and recycle pyridine hydrochloride. Pyridine hydrochloride also chlorinated any magnesium oxide in the spray-dried magnesium chloride. The anhydrous magnesium chloride obtained by the method described contained less than 0.8 percent oxygen. More than 98 percent of the organic amine hydrochloride was recovered.

Jan 1, 1978

By H. Watari

In 1998, a voluntary agreement was reached between the European Commission and the European Car Manufacturers Association (ACEA) to reduce average carbon dioxide emissions of new cars to 25% below 1995 levels by 2008, Agreements have also been reached between the European Commission and the Japanese Automobile Manufacturers Association (JAMA) and the Korean Automobile Manufacturers Association (KAMA), However, the target average of carbon dioxide emissions of less than 140g/km in 2008 is a very challenging goal, Therefore, total product weight reduction is an objective for many manufacturers, This can be achieved utilizing materials possessing high strength and low weight, One material that has recently attracted interest is magnesium alloys, The magnesium alloys are expected to play an important role as next-generation materials, which have possibilities of contributing to lighten the total product weight when magnesium products can be used to replace aluminum and mild steel products.

Jan 1, 2005