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By R. Lasmanis, W. C. Utterback

"The Mount Tolman porphyry Mo-Cu deposit is 64 km south of Republic, Ferry County, Washington. It occurs in the composite Keller Butte pluton, which consists of the Keller Butte porphyritic biotite granite and granodiorite, the Mount Tolman quartz porphyry granite, and the Meadow Creek biotite granite phases. The three phases range in age from 61.3 Ma ± 2.3 Ma to 55.7 Ma ±2.0 Ma. The stockwork deposit occurs in the Keller Butte phase above the contact with the Mount Tolman quartz porphyry granite.Mount Tolman is a hybrid calc-alkaline deposit that has features of both a Climax-type and a plutonic type. Prior to post-ore faulting, the deposit had a flattened hood like shape. In plan it is elongated- 1.6 km wide, 3.2 km long, and up to 609 m thick. The orebody has a 292 o strike. It contains a geologic reserve exceeding 2.18 billion tonnes of ore averaging 0.093% MoS2 and 0.09o/o Cu using a 0.048% MoS2 cutoff. Of this total, 816 million tonnes of measured and indicated ore averaging 0.097% MoS2 and 0.09o/o Cu could be mined by open pit with an overall waste-to-ore ratio of 0.91 to 1.The deposit is a stockwork of quartz and quartz-sericite-pyrite veins and fracture fillings that contain molybdenite, chalcopyrite, sphalerite, and magnetite. Five separate episodes of fracturing and mineralization produced the orebody and were followed by four additional episodes of fracturing with sulrgrade mineralization. The Mount Tolman deposit contains propylitic, argillic, phyllic, topaz greisen, and potassic alteration zones. Extensive silicification overlaps the K-feldspathized, quartz-topaz, and phyllic alteration zones."

Jan 1, 1995

By W. R. Hardwick

Nuclear explosive fracturing has been suggested as a method to make hard rock amenable to extraction by a modified block-caving method. To evaluate this possibility, conventional and nuclear blasting practices are discussed, underground mining methods for large deposits and their costs are described, and the cost of fracturing ore with a nuclear explosive and subsequent mining by a modified caving system is estimated and compared with both open-stope and block-caving mining costs. While the direct cost of fracturing the ore in hard rock with a nuclear explosive is less than with conventional explosives, additional costs imposed by the use of nuclear explosives may make their use unattractive for deep mining.

Jan 1, 1970

By H. R. Butler, R. N. Browne

"The International Nickel Company of Canada Limited consumes some 33 ½ millions of gallons of fresh water and reuses more than 100 millions of gallons of water daily. The continuing search for suitable recycle methodology has enabled the Company to put one 25,000 TPD mill on line utilizing recycled water entirely and a second 35,000 TPD mill is under construction which will likewise employ the concept of recycled water. Future milling plans are aimed at the replacement of all fresh water with recycled water, including sewage effluent and mine waste water.WATER REUSE AT INCO’s SUDBURY MILLSAs man searches unexplored horizons for new and economic industry, he must be constantly and acutely aware of the presence or absence of that precious commodity without which Industry, indeed Man himself, could not survive -- W A T E R. The water requirements of the world today are magnanimous, considering that in 1960 the estimated consumption in the United States was over 300 billion GPD (l) or more than 1,700 GPD per person(5) The magnitude of the problem can be visualized as one considers that three times this quantity of water will be required in the United States by the year 2000, when the dependable supplies of water are estimated to be only two-thirds of that required(2). How, then can we continue to wreak such havoc on that life-sustaining substance - WATER?"

Jan 1, 1970

By A. R. Chambers

Malagash received its name from the Indians many years ago on acco.unt of its turbid (milky) waters, although,' the red men did not appreciate the significance of this turbidity, and were, of course, entirely ignorant of the crystal beds of pure salt that lay beneath their feet. The salt lake of the Mississippian, or Lower Carboniferous period, in which these beds occur, was apparently a large one, for there is evidence of its existence at intervals all the way from Sydney, Cape Breton, to St. John, New Brunswick, 300 miles as the crow flies, and there is no knowing what its other dimensions may have been, as much of this horizon is now buried on the one side beneath the waters of the Gulf of St. Lawrence, and on the other underneath the Atlantic ocean. About midway of the area and underlying the counties of Colchester and Cumberland, there occurs a synclinal basin of which the Malagash salt outcrop forms only a small portion of its northern limit. This basin has been described very fully by A. C. Hayes. (1) The operations carried on at Malagash since Hayes wrote his report have brought to light very little in the way of new geological data, except to confirm his pre-dictions as to the probable faulting of the salt strata, and the effect which these faults would have on mining operations.

Jan 1, 1924

By A. A. Holland

Introduction The world's largest known deposits of natural sodium sulphate (Glauber salt), generally termed 'alkali ' throughout western Canada, are located in the Province of Saskatchewan: reserves are estimated to exceed 200,000,000 tons. These are surface deposits resulting from the recent or present-day natural evaporation of brines, in which respect they are similar to those occurring in certain other parts of the world, as in semi-arid sections of the United States, Russia, Germany and Austria. In some areas, underground deposits formed hundreds of thousands of years ago have been _encountered in drilling, and indications are that they also were formed under arid or semi-arid conditions. The master key to the formation of deposits of natural sodium sulphate, wherever they occur, is a climate, either past or present, under which evaporation exceeds precipitation over a considerable period of time. Commercial development of the Saskatchewan deposits began about thirty years ago. Previously, the demand for sodium sulphate had largely been met by salt cake produced as a by-product in the manufacture of hydrochloric acid from common salt. The rapid increase in the production of Kraft paper and, more recently, the development of synthetic detergent materials, have caused a marked increase in the market for sodium sulphate. This chemical gives the extra toughness to Kraft paper and is used in detergent compositions in proportions as high as 75 per cent. Saskatchewan now has six commercial plants producing natural sodium sulphate. Production from each in the fiscal years 1947 and 1948 is shown in Table I. The newest and latest plant is that of Saskatchewan Minerals, at Chaplin, which 1s designed to produce up to 150,000 tons of salt cake, 98.5 to 99.5 per cent pure, yearly. This is equivalent to about 24 percent of the total United States production of crude commercial salt cake for 1947.

Jan 1, 1949

Morning Star Gold aims to enter the ranks of Australia's gold producers, revitalising the Eastern Victorian Goldfi elds in the second half of 2010 - the fi rst gold production in around 50 years from the Morning Star Gold mine, which was once Australia's largest under WMC. This presentation will outline the redevelopment of the Morning Star Gold Mine from exploration to production. Morning Star Gold NL's (MCO) key project area is centred at the historic mining town of Woods Point and nearby Matlock (to the southeast) and Gaffney's Creek (to the northwest), all of which are located in the lower part of the Victorian Alps roughly 60 kms south of Mansfield. Morning Star has literally numerous historic mines and gold bearing 'dykes' on its tenements, many with a rich production history of high-grade gold from quartz reefs and alluvial sources. The area encompassing the Woods Point goldfi eld and the multiple high-grade gold bearing dyke structures located north towards Jamieson and south towards Walhalla formed one of the most significant mining fields in Australian history. Historical gold production was at least 155 t or around six million ounces of over a century of mining with an average production grade exceeding 27 g per tonne. The bulk of this recorded production came from just a handful of mines and of those only three were mined to any great depth. One of the key historical production hubs in this area was the famous 'Morning Star' mine, which is situated adjacent to Woods Point township. The Morning Star mine produced almost one million ounces at 27 grams per tonne, operating fairly continuously between 1861 and 1959 with a break after Jan 1939 due to the devastating 'Black Friday' bush fi res across eastern Victoria. During the early 1940s, the Morning Star was Australia's biggest gold mine. The Morning Star mine along with Walhalla ensured that this goldfield was third behind Bendigo and Ballarat as Victoria's biggest hardrock gold production centre. Morning Star Gold has rebuilt its 100 per cent owned 'Morning Star' mine with a view to producing gold in the second half of 2010. Morning Star is exploring its regional dykes in order to estimate an overall gold endowment and rank local production sources to feed its proposed mill at the Morning Star mine site. Key initial targets (all within 11 kms of the Morning Star mine) are the Waverly, All Nations, Rose of Denmark, Hunts, Wallaby and Loch Fyne dykes, all of which historically produced high grade gold, much of it gained from only shallow workings above the water table in the weathered zone of rocks. Historically the Morning Star mine was developed over 24 levels to a depth of 940 m via one large vertical shaft and a smaller, deeper inclined shaft at the 1900 feet level underground. At present MCO has dewatered and completely refurbished all infrastructure to the ten level (~310 m below surface). MCO knows there are a number of substantially under-explored and undeveloped or partially developed but not stoped areas within the Morning Star mine. As a result, MCO has been carrying out extensive channel sampling and an underground diamond drilling program targeting areas of the mine that were largely unexplored and undeveloped by the last operators of the mine - Western Mining Corp Ltd (Gold Mines of Australia). Four thousand, fi ve hundred pages of archival WMC records have helped the cause significantly. The shaft refurbishment and reconstruction stage is complete. This has been a program that originally commenced back in 1993 and has cost around $22 M. This refurbishment paved the way for the extensive underground resource focused diamond drilling program, which MCO completed in 2008 with two company-owned and operated drill rigs from drill cuddies on level 9 (900 feet underground) of the mine. This drilling, which encountered some stunning grades intersections and multiple visible gold hits, has outlined further production opportunities including 0.4 m @ 1392 g/t Au (MS 410). Metallurgical testing conducted by Gekko Systems of samples obtained from the exploration program found gravity concentration and a leach circuit gave recoveries of ~95 per cent to 99 per cent, or ~92 per cent from gravity alone at a coarse crush (300 ¦m). Currently, the plan is to use gravity to fi rstly produce dore bars on-site from the free milling gold and secondly a saleable concentrate for CIL processing off-site. A purpose built 10 t/h (~80 000 t/a) gravity treatment plant was subsequently ordered from Gekko and is expected to be installed onsite and commissioned in the Q3 2010. The plant features fi ne crushing with a vertical shaft impactor, followed by gold concentration by inline pressure jig, inline spinner and Wilfley table. This is an ABSTRACT ONLY. No paper was prepared for this presentation.

Aug 1, 2010

By F. Göktepe

Flotation gases and conditioning of the pulp with them are one of the important parameters of flotation and should be taken into account especially in the flotation of sulphide ores. In this study, flotation of pure chalcopyrite and pyrite samples were investigated with air, oxygen and nitrogen as flotation gas and the effect of conditioning with these three gases also were performed. Flotation gas was introduced for 5min, 10min, 15min and 20 min. For chalcopyrite flotation there was no drastic change in recovery when air, oxygen and nitrogen were used as flotation gas but conditioning the pulp with oxygen for 10 minutes increased the recovery about 8% where pulp potential has been approximately 25 mV with chalcopyrite and 150 mV with platinum electrodes. Conditioning with all three gases decreased the pyrite recovery by about 20% with 20 minutes conditioning. Pulp potential decreased in the presence of nitrogen while conditioning with air and oxygen increased the pulp potential for both electrodes. Generally potential increased for both minerals with oxygenation and aeration time but it was stable or slightly decreased with nitrogenation. Comparison of the gases showed that the different flotation gases do not effect flotation recovery but conditioning the pulp with these gases at pre-flotation effects chalcopyrite and pyrite flotation recovery, most dramatically pyrite.

Jan 1, 2014

By David Mast, Irwan Halim, Seung Han Kim

Nowadays more infrastructure tunnel projects are procured using design-build method. In this environment, the Owner’s engineer would typically provide performance requirements for the design-builder, and the Contractor’s designer would work with the contractor to design the tunnel lining based on the Contractor’s means and methods and manufacturing process. This paper will describe the typical requirements provided by the Owner based on its design objectives, and the means and methods by the Contractor that would significantly impact the design. Project examples and recommendations for Owner’s requirements will be provided that can optimize the design and still achieve the Owner’s performance objectives.

Jun 13, 2021

By E. P. Hayes

This report supplements the date contained in Bureau of Mines Information Circular 6156, "Special features of core drilling in the salt beds of Western Texas and New Mexico," by J. S. Wroth. That circular dealt with the drilling of the first 12 of the 24 potash tests made by the U. S. Bureau of Mines in cooperation with the U. S. Geological Survey; this report is concerned with the drilling of the last 12 tests. ACKNOWLEDGMENTS The writer is indebted to R. E. Heithecker, the Bureau of Mines engineer who had charge of the drilling of tests 13 to 17, for his well-prepared records on the drilling of these tests.

Jan 1, 1933

By Eric E. Cather

A mineral survey of a portion of the Lake Eleanor RARE II area (No. 5807), Trinity County, California, was done by the U.S. Bureau of Mines and U.S. Geological Survey (USGS). Approximately 8,650 acres (3,500 ha) of the Lake Eleanor RARE II area not examined for this report had been previously investigated by U.S. Bureau of Mines and U.S. Geological Survey personnel and published as part of USGS Bulletin 1371-B, (Hotz, 1972) Mineral Resources of the Salmon-Trinity Alps Primitive Area, California. The report showed that the section included in the rake Eleanor RARE II area contained no metallic mineral resources. Geochemical sampling indicated that the Lake Eleanor RARE II area contains no potentially important mineral resources. About 4,650 acres (1,880 ha) of the eastern part of Lake Eleanor RARE II area were examined for this report. The Bureau of Mines was responsible for detailed examination and sampling of mines, prospects, and mineralized zones; the Geological Survey conducted geological, geochemical, and geophysical studies of the area. This area is underlain mainly by granitic rock and serpentinite.

Jan 1, 1981

By R. Viljoen, K. Manns

"The mining industry instinctively follows rule-of-thumb design velocities often based on very simplistic theoretical “economic velocity” optimization calculations to determine airway sizes. Although these are important building blocks in ventilation planning it should be recognized that mines are far more complex and various, often opposing factors, are at play. Therefore, it is important that each ventilation situation be analyzed on a case-by-case basis to determine the best-fit solution for an operation. The ventilation design process should not necessarily be constrained by theoretical optimal economic velocities that become rules-of-thumb. In some cases, this could lead to unrealistic infrastructure requirements and exorbitant capital expenditure. This type of design often lacks corporate support leading to a scenario where no ventilation upgrades are implemented. Economic velocities typically consider the cost of power, development costs and operational risk. These typical rules-of-thumb work reasonably well as a starting off point when considering feasibility of greenfield projects. For established mines that are looking at expansion these rules may not be appropriate due to capital expenditure constraints or geotechnical ground conditions. The theoretical optimum should be reviewed in the case of an established mine where existing infrastructure could perhaps be maximized. For example, the geotechnical risk of large diameter raises at depth may require multiple small diameter raise to maintain the theoretical economic velocity. If a mine has a life of less than 5 years, then the cost of power is an insignificant proportion of the total owning cost. For these mines, operating the ventilation raises at an increased velocity to avoid a complicated and expensive capital upgrade might be the more appropriate option, even though it opposes the traditional rules-of-thumb. Mines which are capital constrained may be better off accepting a higher operating cost for the interim or settle for minor infrastructure upgrades rather than run the risk of the entire ventilation upgrade not being approved at all. This paper explores a number of examples derived from case studies undertaken by the authors for several deep mechanized mining projects in Australasia. The examples explore conditions under which higher than the economic norm would be applicable."

Jan 1, 2019

By Tyler Blake

This report describes a study designed to evaluate a pro-posed modification in the cap lamp battery cord intented to reduce the incidence of snagging and catching of the cord. A three phase evaluation was performed comparing a prototype with the standard, currently available cord. First, a ?pull test? equipment was run. The second evaluation was a body conformity study. The third evaluation involved subjects performing common tasks in a low seam coal mine simulator. The results of these evaluations clearly demonstrated significant safety advantages of incorporating coiled cord into the design of the cap cord. The coil cord presented less of a snagging hazard, allowed the wearer more time to respond to a snag, and transmitted lower levels of force to the helmet when snagged. The research team recommended that cap lamp cords be commercially developed which incorporate coiled cord at the battery and helmet.

Jan 1, 1980

By Ahmad-Bitar¸ Riyad N., Shadia J. Ikhmayies

"CdSi-xTex thin films with O<x<0.3 were prepared on glass substrates by the annealing of CdS:In thin films produced by the spray pyrolysis technique. The annealing was performed in nitrogen atmosphere at 400 --,C in the presence of Te vapor. The compositions of the films were determined by energy dispersive X-ray detection (EDAX) measurements. Scanning electron microscopy (SEM) was used to investigate the morphology of the films. The transmittance of the films was measured at room temperature in the wavelength range 400-1100 nm and used to produce the absorbance of the films. The first derivative of the absorbance was used to estimate the optical bandgap energy Eg by using the minima in the first derivative. The relation between the bandgap energy and the fraction x was found to fit with a parabola and bandgaps in the range 2.196: =Eg: =2.480 e V were obtained. Urbach tailing in the bandgap was also investigated.IntroductionPolycrystalline thin film CdS/CdTe solar cells are promising candidates for low cost and high efficiency photovoltaic applications. The fabrication process of these cells involves the deposition of a p-type CdTe absorber layer on top of an n-type CdS one which acts as a window. The efficiency of the solar cell is affected by the method of preparation of both layers [1, 2].Interdiffusion processes taking place at the CdS/CdTe interface have largely been realized to significantly affect the device performance [3] by changing the bandgap of both CdTe and CdS which modifies the spectral response of the solar cell [ 4]. The formation of a mixed crystal or a graded layer causes a reduction in the lattice mismatch at the CdS/CdTe junction [5] which is about 9% [ 6] and then a reduction in the number of interfacial states and recombination centers, thus enhancing the solar cell performance [7, 8]. A complete understanding of the properties of the intermixed layers CdSxTe1-x that usually constitutes the interfacial region in the CdS/CdTe solar cell is essential. According to our knowledge, a lot of research was devoted to the investigation of the diffusion of S into CdTe [3-5, 9]; but little on the direct investigation of the diffusion of Te into CdS [6]."

Jan 1, 2011

By M. Ahrens

The paper describes the main drive systems of an iron ore pelletizing plant and related important design aspects. It is shown how today?s drive systems provide the desired flexibility for the process and lead to high energy efficiency and system reliability. With respect to electric drives expansion projects have many technical issues in common with green-field installations. However, there are aspects where important, distinct differences occur. It is shown which considerations for selecting and sizing drive systems for such a plant as well as for integrating them into an overall plant control concept need to be taken into account. Besides the design of the drive systems a key aspect for such an expansion project is the network interaction. Harmonics levels at different points in the plant network need to be checked carefully with simulations during the design phase to avoid problems later-on during operation. The power factor compensation and the harmonic filtering of the new production line cannot be designed independently of the already existing plant. The paper highlights areas where problems could occur and which points need to be paid attention to when a plant is expanded.

Jan 1, 2008

G. D. Kneip, Jr., and J. 0. Betterton, Jr. (Union Carbide & Carbon Corp., Oak Ridge, Tenn.)—The authors have contributed to the theory of zone melting by considering the effects of the solidification of the final zone on the distribution curves for the finite length bar. However, they do not consider in sufficient detail the restrictions imposed by the particular metals involved. Eq. 4 has the analytical solution which is equivalent to Pfann&apos;s&apos; equation for normal solidification of the final zone. For distribution coefficients less than one, it is apparent from this form of the equation that, as the solidification approaches completion, the concentration attains extremely large values which are inconsistent with the density of physical materials. The necessary restrictions then are that the solute concentration cannot exceed the density of the alloy, or more frequently, the solubility limit in the solid phase. Secondly, the distribution coefficient cannot be constant from 0 to 100 pet solute unless the liquidus and solidus coincide over this whole region. It would thus be more correct to impose limits on the solubility such as would be indicated by a typical eutectic diagram. Furthermore, in this case the assumption that the distribution coefficient remains constant is more likely to be realized. Eq. 4 should then be replaced by the following C,,(x) =k/L-x [ ?11 3 C11 (x)Ddx - ?11 4 Cn(x)dx= k[Cn (L-l)/k] [ L-x/l]x-1 L-l = x = L-l [kCenterlie/Cn (L-l)]1/k-1 The typical shape for the distribution curve in the final zone is indicated schematically in Fig. 12 for the simple eutectic case. The solidification proceeds according to Pfann&apos;s&apos; expression until the concentration reaches the maximum solubility in the solid phasc. At this point, the concentration changes abruptly to the eutectic composition and remains at this value for the duration of the solidification. The width of the flat region, and the back reflection of this effect into the concentration curve on subsequent passes, depends upon the ratio of the eutectic concentration to the original concentration, upon the distribution coefficient, and upon the number of passes. Similar effects on the concentration curves in the first zone length would be expected in many systems for which the distribution coefficient is greater than one, and where the maximum solubility is not too much larger than the original concentration. The discussers agree with the authors that the assumption that the liquid is uniform in concentration should be considered with caution. Kneip and Better-ton&apos; have shown, however, that in floating zone refining of zirconium, using induction heating, the distribution of iron agrees with theory which assumes a uniform liquid composition. Hence, in this case, the experimentally realized distribution coefficient agrees with the present phase diagram within the experimental uncertainty. L. Burris, Jr., C. H. Stockman, and I. G. Dillon (authors&apos; reply)—The authors are happy to have received the comments of G. D. Kneip, Jr., and J. 0. Betterton, Jr., which provide additional insight into the zone refining process. It is true that the equations developed are inapplicable if the solute solubility in the solid phase is exceeded. However, this restriction on the use of the equations was clearly stated in the paper. A closer approach to

Jan 1, 1957

By Chen Xia, G. Deschenes, Wan-Tai Yen

"Ammonium thiosulphate is a potential replacement for the cyanide in gold leaching. However, the high reagent consumption hinders its practical applications. In order to reduce the thiosulphate consumption, chemical additives were used. The gold leaching tests were conducted on a sulphide copper-bearing gold ore with 0.3M thiosulphate, 0.03M copper sulphate and 3M ammonium hydroxide solution at 44% solids and pH 10.2. Sulphate, sulphite, amino acid and chelating agents were initially used as additives. It was found that chelating agents (EDTA and NTA) could reduce the thiosulphate consumption from 30 kg/t to 17.8 kg/t with some increase of gold extraction. The improvement from other additives was not significant. Other approaches were also attempted to reduce thiosulphate consumption with some success. The thiosulphate consumption could be reduced to about 17 kg/t by reducing copper sulphate concentration fiom 0.03M to 0.0015M; adding 0.04M of EDTA; replacing copper sulphate with 0.001M of nickel sulphate. The increase of pulp density fkom 44% to 60% solids could reduce the thiosulphate consumption to 17 kg/t as well. Limiting the dissolved oxygen content by sparging nitrogen reduced the thiosulphate consumption further to 10 kg/t. It could be concluded that the regulations of copper and oxygen supply in the lixiviant were the keys to reduce the reagent consumption.INTRODUCTIONCyanidation, as the predominant gold leaching method, has been used in mining industry for more than a century. The low reagent requirement and the high performance are the reasons for its great success. However, cyanidation cannot deal with ores containing large amount of copper, nickel, arsenic and zinc minerals. This type of ore may consume more reagents and result in a poor gold extraction. On the other hand, the carbonaceous materials may preg-rob the gold from pregnant solution and make cyanidation a problem leaching method. Further more, the high toxicity of cyanide represents another challenge to modern gold mines. In the beginning of the year 2000, an environmental disaster caused by an accidental spill in Yugoslavia was broadcasted world-widely (Gurevitch 2000). Gold mine operators are facing higher social pressures and increasing costs on environmental protections. The usage of cyanide in mining industry has been banned in some countries and locations such as Czech Republic, Japan, Turkey, California, Colorado, Wisconsin and Montana States. Thus, a non-cyanide lixiviant is required."

Jan 1, 2002

By Jean-Christophe Ehrström, Fanny Mas

"A specific test set-up has been developed to study at the lab-scale the high-temperature behavior of aluminium alloys in conditions close to those experienced by thick plates during hot-rolling. A particular attention is paid to reproducing both the stress triaxiality and strain fields that exist in plate during rolling. In addition to standard ductility measurements by means of secondary tensile tests (STT), multi-step plane strain compression (PSC) tests have been performed to mimic hot-rolling pass schedules. Tool and sample geometries were adapted to allow low h/a ratios (h= average specimen thickness, a= tool width) and thus guarantee depressive stress conditions at the center of the specimen during most of the test. Such conditions resulting in porosity opening can then be used to investigate the impact of rolling conditions on damage evolution. The mechanical testing was combined with post-mortem high-resolution ultrasonic scans allowing to detect small amounts of damage and quantify its evolution as a function of processing parameters such as strain and temperature. The test set-up was validated on two 2050-type AlCuLi alloys with different susceptibilities to high temperature embrittlement.INTRODUCTION Aluminium thick plates are widely used in the aerospace industry, mainly as semi-products to be machined for internal structure parts of wings (spars, ribs) and fuselage (frame). The conversion route of such products always involves a hot-rolling step, the conditions of which (temperature, speed, pass schedule) have to be optimized to maximize both mechanical properties and final plate quality. This goal cannot always be efficiently achieved by full-scale industrial trials as (a) cost limits the number of possible test conditions and (b) it is often difficult to change one parameter alone. That is why there is a need for laboratory-scale tests able to mimic the rolling process in terms of temperature, strain rate, stress triaxiality and deformation. Most laboratory hot-rolling mills do not adequately simulate industrial processing routes, especially in terms of heat loss, deformation speed and rolling geometry (maximum roll gap vs. work roll diameter). Concerning thick plate, it is of interest to achieve both similar deformation and L/h ratio (L=arc of contact, h=mean gauge) per pass as the industrial schedule because they are simultaneously involved in damage evolution and void closure during hot forming processes (Saby, 2014)."

Jan 1, 2018

By O. Jerez, A. Novoselov, L. Gutiérrez, J. Morales, U. Kelm

"In Chile, the evaluation to reprocess tailings for elements other than the traditional Cu, Mo, Au, and Ag, as well as projects to implement long-term monitoring of mining residue sites are key elements of the government’s roadmap laid out for the mining industry for the next two decades. Characterization routines for tailings and mining residues in general have to be adapted from traditional ore and gangue mineral and micro-chemical studies. This contribution summarizes key aspects for the regional mining industry for a basic characterization of tailings with emphasis on small- to medium-size operations with limited in-house facilities and funds to carry out extensive characterization programs. The oldest tailings in Chile were deposited almost a century ago; this fact gives importance to the recovery of past ore /gangue feed information, process circuits and changes (e.g. acid to alkaline flotation), process water provenance and recovery, meteorological registers, and routines of filling a tailings impoundment. Basic characterization should include X-ray fluorescence spectrometry combined with laser diffraction (LD) granulometry followed by X-ray diffraction (XRD), optical microscopy (OM) and automated mineral analysis (AMA) on a reduced set of samples identified by hierarchical cluster analysis. For this detailed stage, only coarse tailings are recommended for direct a study by microscopy and automated mineralogy (P80 >100 µm); for fine tailings (P80 ~30 µm) a combination of XRD and deposit type adapted sequential extraction are more informative options. In particular, for XRD of clay phases, losses of crystallinity may result in deviation of spectra from standard reference phases due to prolonged storage. To monitor surface alteration, scanning electron microscopy can be complemented by Raman spectroscopy to detect residual reagents on mineral surfaces. For reactivity modeling BET measurements are essential."

Jan 1, 2018

By Staff

This paper contains an outline of the history of the Sullivan mine, now owned and operated by the Consolidated Mining & Smelting Company of Canada, Ltd.; some account of the various methods by which the ore has been treated in efforts to solve the metallurgical problems which its complex character presents, and, finally, to give, as concisely as possible, an outline of the present mining and metallurgical practice along with a description of the mining and milling plant at Kimberley and the smelter and refineries at Trail, in British Columbia. EARLY HISTORY In May, 1892, Pat Sullivan, John Cleaver and Mike Holland left the Coeur d&apos;Alene country, in the State of Idaho, to seek their fortunes in the Kootenay lake district in British Columbia. They were aided in their venture by James Cronin, afterwards locator and owner of the St. Eugene mine at Moyie. Two months later the party broke up, Sullivan and Cleaver crossing the mountains from Crawford bay on Kootenay lake to the head waters of the St. Mary&apos;s river, the course of which they followed down to Fort Steele. Here they found a good deal of excitement about the North Star mine, and, joined by "Ed." Smith, who was familiar with this part of the country, they determined to return and prospect in its vicinity.

Jan 1, 1924

By A. C. Smith

Although only used in a few U.S. longwall mines, bleeder-less ventilation systems can be an effective spontaneous combustion control method in mines having a demonstrated history of spontaneous combustion. To provide insights for the optimization of bleederless ventilation systems for U.S. underground coal mines, a computational fluid dynamics (CFD) study was conducted to model the spontaneous heating in longwall gob areas using a bleederless ventilation system. A single longwall panel with a bleederless ventilation system was simulated, and typical longwall mine ventilation data were used in the simulations. The permeability and porosity profiles for the longwall gob were estimated using a geotechnical model and were used as inputs for the CFD modeling. The effects of gob permeability and resistance of the collapsed entries on the spontaneous heating were studied. The effectiveness of using nitrogen injection to prevent spontaneous heating in the gob was also examined.1

Jan 1, 2008