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A Seismic Hazard Scale (SHS) has been developed to provide a practical seismic hazard assessment tool applicable to mine site personnel. It has been applied to seismic data sets from numerous Western Australian mines. The SHS is applicable when quantifying seismic hazard on a mine wide scale, on a mining block scale, and on a local cluster-by-cluster scale. Spatial mapping of variations in seismic hazard is a particularly powerful application of the scale. The SHS is sensitive to temporal increases in seismic hazard. It has also been shown that the SHS can be applied to estimate seismic hazard when seismic monitoring data are not available. The SHS is a simple derivation of the Gutenberg-Richter relation. There are assumptions and limitations to the SHS that need to be considered when applying it to mine seismicity data.

Jan 1, 2004

By M. Mostert

Mining companies sponsor a range of non-core, corporate social responsibility projects to adhere to social and labour plans and environmental management prerequisites that form part of a mining licence application. Some companies go above and beyond such projects, sponsoring initiatives that generate renewable energy through solar power, wind energy, natural gas, etc. The challenge for these companies is to choose between a variety of projects to ensure maximum value, especially in times when the economic climate might be less favourable for such projects. The focus of this research was to analyse the concept of sustainability as it exists today, and to apply that to the triple bottom line accounting method in an attempt to quantify the sustainability of a project. Research was done on the methane burn-off project at Sibanye Gold?s Beatrix Mine to establish how such projects are planned and financed, and what impact they have on the triple bottom line of a company. The financial bottom line is, by definition, one that executives understand. This paper also proposes a quantitative method for defining the social and environmental bottom lines as well. By considering the financial, social, and environmental values, the study attempts to determine a monetary value for a sustainable renewable energy project. This monetary value can be compared to similar values obtained for other sustainable renewable energy projects under consideration. The research suggests that monetary value alone is not enough to base a sustainable decision on, and qualitative measures are suggested for use in conjunction with quantitative methods. The selection method proposed should assist the board of a mining company to choose the most sustainable option and the project that will add the greatest value to the company across all three bottom lines. It will also provide increased justification for such renewable energy projects, even in periods of harsh or uncertain economic climates.

Jan 1, 2014

By F. D. Wright, H. D. Fara

The solution to the plastic and elastic stresses around a circular shaft in a hydrostatic stress field that has appeared in two editions of a widely used book on rock mechanics1 is in error. Although this is a relatively minor matter, the correct solution was not found by a literature survey, so it appears worthwhile to present it. For a homogeneous, isotropic, rigid plastic-elastic body in plane strain which obeys Tresca's yield criterion the following equations must be satisfied:'," In the plastic zone,

Jan 1, 1963

The National Mining Hall of Fame will induct five mining greats on Saturday, Sept. 17, 2005 at the Museum Convention Center in Leadville, CO. The 18thAnnual National Mining Museum and Hall of Fame Induction Banquet will begin with an open house and reception at 5:30 p.m., followed by the dinner and program at 7 p.m. The five inductees are Nathaniel Peter Hill, John Munro Longyear, John McDonald, Julius Eckhardt Raht and John Dennis Ryan. Nathaniel Peter Hill (1832-1900) built the first successful smelting operation in the Rocky Mountains in1867 at Black Hawk, CO. From 1879 to 1885, he was a U.S. senator. He also taught courses at the Colorado School of Mines.

Jan 1, 2005

By André de Albuquerque Vicente

Foram estudadas as microestruturas de quatro juntas soldadas de aço inoxidável austenítico produzidas através do processo MIG/MAG com diferentes gases de proteção, dentro de um projeto que visava estudar o comportamento de juntas soldadas de aços inoxidáveis austeníticos a altas temperaturas. Utilizou-se o metal de adição 22.12.HT e diferentes gases de proteção: 100% Ar, Ar + 2% CO2, Ar +4% CO2 e Ar + 20% CO2. Para cada mistura gasosa, foi soldado um corpo de prova. Estudouse a variação das frações volumétricas de ferrita delta em função dos diferentes gases de proteção utilizados. Observou-se aumento no teor de carbono e consequente diminuição das frações volumétricas de ferrita delta nos metais depositados com gases de proteção com maiores concentrações de CO2, evidenciando-se que o carbono é um forte estabilizador da austenita nos aços inoxidáveis austeníticos.

Aug 17, 2017

Outotec?s direct leaching process is an atmospheric leaching process for sulphidic zinc concentrates. The direct leaching process, based on special leaching reactors developed by Outokumpu, is the most economical way to produce electrolytic zinc. The production scale applications of the process realized so far are in use at Boliden?s Kokkola and Odda zinc plants. A new full-scale application of Outotec?s direct leaching process will be started in 2008 in China. The process is integrated to the existing production plant, replacing part of the old production facilities by modern technology. Besides high zinc extraction the plant expansion and modernization offers also indium recovery, iron precipitation and solution purification processes. The process equipment includes tailor-made reactors designed on the basis of the work and experience of Outotec research center. Keywords: zinc, concentrate, atmospheric leaching, technology, indium, iron precipitation, solution purification, Outotec

Jan 1, 2008

By J. J. Jakosky

THE method of electrical mapping of oil structures to be described possesses certain limitations, as well as certain definite advantages. It, in common with other geophysical methods, is not a panacea for the solution of all subsurface geological problems. It operates on differences in electrical conductivity of the geological strata, whereas the other most-used methods operate on (1) differences in magnetic properties, (2) differences in gravimetric properties, or (3) differences in seismic properties. This modification of the resistivity method is therefore able to obtain subsurface information under conditions not favorable to the other methods mentioned. Where the other methods have been employed with ambiguous results it may act as a complementary method.

Jan 1, 1936

By A. P. I. Popoola, D. O. Okanigbe

Copper smelter dust (CSD) is a byproduct of copper smelting and is composed of considerable amounts of copper values and also contaminants (like silica and mullite). Silica in CSD forms Siloxane, a low density material, which is difficult to separate from the aqueous phase during solvent extraction. The need therefore arises to reduce/remove the aluminosilicates for better grade and recovery of copper in subsequent processes. This paper reports the use of two types of centrifugal gravity concentrators, namely Knelson (KGC) and Falcon (FGC), for upgrade of copper in CSD. The aim was achieved by an initial characterisation (Particle Size Distribution (PSD) and mineralogical analysis), together with Density Separation (DS) experimentation, premised on a hypothetical density difference between copper oxide and silicon/aluminum oxide. Wet screening results showed that 88.52 % of CSD particles were in -53 μm fraction. Further PSD study using laser diffraction showed the -53 μm fraction to be composed mainly of 26 μm, 22 μm and 18.5 μm fractions. Mineralogical analysis showed copper to occur mainly as oxides (24.34 wt%), but also as sulphides (7.95 wt%), and a substantial amount of aluminosilicates (42.97 wt%) were identified. Density of CSD was determined as 2.83 g/cm3. Copper content (18.02 wt%) in CSD was upgraded to 35.94 wt% in concentrate using KGC. In conclusion, the results showed good convergence between hypothesis and experiment. Additionally, it was also concluded that KGC performed better than FGC for copper upgrade in CSD. However, roasting CSD prior density separation was recommended for enhanced copper upgrade. Keywords: Copper smelter dust, density separation, solvent extraction, mineralogy

Jan 1, 2020

By Rowan A' Murry Karstel

One way to ensure dragline productivity is by Increasing the size of the dragline bucket. In August 1994, Optimum Colliery in conjunction with Van Reenen Steel, Northwest Applied Technology, SSAB of Sweden and Barlows Equipment launched the new lightweight dragline buckets. Optimum built the first prototype bucket that was used for the deslgn for future buckets. After knowledge had been gained from measuring the input force of the bucket Northwest Technologies designed the new bucket with finite element analysis. Together with the knowledge of quenched and tempered steel from SSAB of Sweden and Barlow Equipment, the new Wrightech bucket was born.

Jan 1, 1998

By David R. Beadman

The Fleming method is widely used for the analysis and prediction of single pile behaviour under maintained loading. The method does not distinguish between bored and driven piles. Driven piling avoids the need to dispose of spoil from the pile installation and is therefore becoming increasingly attractive. This paper considers the theoretical effects of the possible preload at the base of a driven pile after installation, using the hyperbolic functions as defined by Fleming. The Fleming method is reviewed and the equations and analyses are adapted, using spreadsheets, to allow for a possible preload. The simplifying assumption of the effect of the elastic compression of the pile is also reviewed and a more analytical method adopted to assess the elastic compression. Some driven pile test data is analysed to consider the possible effects of the preload on the performance of a driven pile and to test the theoretical relationships that are developed in the paper.

Jan 1, 2006

By Paul D. Tomlingson

Part two of this two-part series covers the use of an integrated information system with administrative information, including error and absentee reports as well as open work files. Part one (ME, July, page 650) discussed conceptualizing and developing a maintenance information system. It explained what information is needed to manage such a system.

Jan 8, 1985

Open pit mining has considerable damaging effects on the environment of the mining region. These harmful influences have to be controlled during mining operation until mine closure. Different sections of mined land, such as; pit(s), waste dump(s), tailing pond(s), roads, areas for onsite facilities and free land zones which are not mined, have to be reclaimed through treatment of environmental pollution and implementation of a permanent post mining land use (PMLU) after mine closure. Implementation of a permanent optimal post mining land use (OPMLU) is one the most decisive criterion which effects on reclamation procedure, mine closure and their costs. Therefore to define ultimate pit limit (UPL), mine closure and reclamation costs should be entered in the calculation iterations same as the other variable costs of the projects. Estimation of closure and reclamation costs of a mining project is carried out based on the attained OPMLU and the specifications of the mine site. In this paper, a model is developed to provide OPMLU for pit area through fuzzy multi-attribute decision making (FMADM) technique. The model was implemented on a hypothetical copper mine in northwest of Iran as a case study. Lumber production was selected as OPMLU for the pit area. It is finally concluded that through the developed model, OPMLU for different sections of mined land in an open pit mine project is defined. Also closure and reclamation costs are estimated based on the defined OPMLU to achieve a more accurate UPL.

Dec 6, 2010

By H. Kinilakodi

In 2008, the authors developed a methodology for assessing underground coal mines for high risk for major-hazard events. It focused on major hazard-related violations of safety standards associated with high-risk conditions. Later using a sample of 103 mines with 2008 and 2009 data, injury measures and MSHA citation data were combined into a Safe Performance Index (SPI). The SPI was used to assess the relative safety-related risk of mines, including by mine-size category. In this paper, comparisons are made of SPI-associated, normalized safety performance measures across mine-size categories, with emphasis on small-mine safety performances as compared to large and medium-sized mines.

Jan 1, 2011

By Jörg Holzhäuser, Matthew Crow

An 18.5 km (12 mile) long tunnel is being driven by four Earth Pressure Balance Tunnel Boring Machines (EPB-TBMs) of 4.7 m (15.5 ft.) diameter through alluvial sand, gravel, silt and clay above and below the groundwater table in Los Angeles beneath buildings, railroads, freeways and the L.A. River. Continuous automatic data recording of TBM operation and field inspectors' records, accompanied by monitoring of 5,000 surface reference points and 268 borehole extensometers were used to obtain information. The paper explains the reason for selecting the method of EPB-tunneling to ensure stability of the tunnel and meeting of environmental requirements, and presents data collected to monitor and demonstrate the performance of the TBM sin the different ground conditions encountered.

Jan 1, 2003

By C. C. Miller, A. R. Belyea, C. H. Herty, E. B. Burkart

The removal of sulfur from steel has been studied by many investigators, but the quantitative relationships between the factors involved have not been determined. This is undoubtedly due to the number of variables encountered in practice, an attempt to study any one reaction or relation being affected by other reactions the relative importance of which cannot be estimated. The factors that affect the extent of removal of sulfur in the basic open-hearth process are generally considered to be: 1. Basicity of the slag: (a) Iron oxides in the slag; (b) carbon in the metal or carbide in the slag; (c) temperature. These factors affect the reactions: FeS + CaO = CaS + FeO FeS + CaO + C = CaS + CO + Fe. 2. Amount of sulfur in the gas. 3. Amount of manganese in the metal. The factors which affect the rate of removal of sulfur include, in addition to these: 4. Fluidity of the slag: (a) Normal changes in fluidity with temperature; (b) changes in fluidity because of the addition of special reagents such as fluorspar. 5. Agitation of the bath. Within the past two years, two papers have been presented dealing with the influence of sulfur in the gas on slag and metal. Nead' has shown that if a gas high in sulfur be used for heating, the metal and slag will absorb sulfur. Whitely2 concluded from his experiments that if a

Jan 1, 1925

By M. Dobbins

The advent of the more efficient and effective rare-earth roll (RER) separator in the early 1980s began a new era of dry magnetic separation. The RER units quickly became the separator of choice within the materials processing industry. Though RER separators are now commonplace throughout the world, only modest advancements have been made to the original design. This paper details recent, more significant, rare-earth magnetic separation developments. These developments maintain the best features of existing technology while minimizing shortcomings. The discussion includes discerning features that aid in better metallurgical performance. Also discussed are improved mechanical aspects that allow for both easier operation and maintenance. Performance details from various test programs are included.

Jan 1, 2010

By K. A. Grace

K.A. Grace's article advocates usage of the terms describing resources as defined by the US Bureau of Mines (USBM) and the US Geological Survey (USGS). Originally recommended in 1943, the terms were revised in 1976 and again in 1980. Mr. Grace further recommends that the term reserves should be severely restricted and that it should encompass mining and processing factors. The list of references quoted in the article does not, however, mention the study concerning reserves that was made over a period of almost four years by a joint Committee on Definitions. This committee was sponsored by three societies - the American Institute of Mining, Metallurgical, and Petroleum Engineers (AIME), the Society of Economic Geologists (SEG), and the American Institute of Professional Geologists (AIPG). Most of the detailed work was done by a working party appointed by the Society of Mining Engineers (SME) of AIME. After studying previous definitions in the US as well as in the United Kingdom, Canada, Australia, France, and other European countries, the following definitions were proposed and accepted in 1974 by the SME board of directors. (Ore reserve definitions in Rhodesia, South Africa, or New Zealand were not mentioned as having been reviewed.) Table 1 lists definitions from MINING ENGINEERING, July 1975. Notice in Table 1 it was mentioned, under Resources, that the definitions of reserves were recommended for the purposes of commercial evaluation. It was further stated that the terminology adopted by the USBM and the USGS, as released to the public, covering resources, was recognized as such. Mr. Grace also recommends that reserve calculations should allow for future losses in ore tonnage, ore grade, milling, and smelting. Unfortunately, it might take years to develop a prospect after it has been discovered. Mining methods have to be tested before being finalized in order to know what tonnage of ore is going to be lost, or how much dilution is likely to occur. The same problems are present in concentrating or in smelting. Operating mines usually know how mined grade compares with ore reserve grade. They can usually predict fairly closely what their final metal recoveries are probably going to be. It is, however, expecting too much to have to delay quoting reserves from say a recently drilled prospect. Any responsible mining geologist or mining engineer (or company for that matter) should state several things about an ore reserve calculation: that it is or is not a preliminary estimate only; that estimated grades have not been diluted to allow for a future, as yet unknown, mining method; that metal recoveries are or are not only tentative, based perhaps on drill hole samples only or on no metallurgical testing at all. Reserves of a partially explored prospect are always quoted when it becomes necessary to raise capital. If they are described as preliminary only (really just tonnage and grade estimates and not true "reserves") in addition to being accompanied by some of the items just mentioned, nobody is going to be misled. If, on the other hand, a report by a member of the SME or SEG is intentionally misleading, the individual must be reported and eventually disbarred from his or her society. To summarize, let us stay with the definitions of ore, ore reserves, and related terms as recommended and accepted by SME. I might add that on occasion I have used a fourth category in my ore reserve calculations, namely Geologically Possible Ore. This category is more or less the same as Inferred ore defined by the USBM and USGS. I have never given a grade to this category, which may be based only on geological mapping, geochemical sampling, or geophysical traversing.

Jan 8, 1985

By M. Rebolledo, N. Mora, G. Zarate

For decades, the mining industry has been focussed to find an efficient and economically attractive process for the leaching of chalcopyrite. Jetti Resources has developed a novel hydrometallurgical alternative to extract metals of value from mineral ores that the mining industry has struggled to process. Jetti’s catalytic technology, whilst being especially advantaged for extracting copper from low-grade and highly refractory minerals, such as chalcopyrite, seamlessly integrates with existing heap leaching methods and downstream processing operations. The technology uses a catalyst which weakens the metal-sulfide bond in the crystalline structure of the chalcopyrite mineral and alters the structure of the passivation layer allowing the leaching process to continue. Metallurgical test work was conducted to examine the effect of ore variability, in terms of grade and copper deportment. The results presented in this study highlight the amenability of the technology and demonstrate that it is possible to enhance leaching performance of copper primary sulfides over a wide range of copper grades, from 0.1 to 2.2%. Further, it is concluded that despite the copper grade, the tests with increasing chalcopyrite content render copper recoveries up to 4 times higher compared to the tests under non-catalyzed conditions. Jetti’s catalytic technology is currently being evaluated at larger scales, namely a commercial dump leach operation and 50-tonne crib tests. The results obtained thus far are compared to the results from conventional sulfide heap leaching tests. INTRODUCTION For decades, the mining industry has been focussed to find an efficient and economically attractive process for the leaching of chalcopyrite. Chalcopyrite (CuFeS2) is the most abundant and recalcitrant of the copper sulfides (Singer, 2017). Owing to its abundance and economic importance, it is the most demandable natural source for copper worldwide. In fact, the projected demand for copper is further expected to rise much higher beyond the current demand by the year 2020 (Panda et al., 2015). Therefore, it becomes imperative to find low intensity hydrometallurgical technologies for the extraction of copper from chalcopyrite in ores of such low-grades that are un-economic to concentrate and to augment the supply by extracting copper from ores with complex mineralogy that cannot be concentrated (Watling, 2013).

Jan 1, 2019

By T. R. Raponi

"The Williams Mine, Canada's largest gold mine, processes over 6,000 t/d of ore utilizing SAG/ball milling, leaching and CIP, carbon stripping, electrowinning and refining to produce gold dare bullion.Grinding accounts for over 50% of milling costs with grinding media, power and mill liners making up over 75% of grinding costs. Since 1991, several changes in these areas have been made which have significantly reduced costs, while fineness of grind and mill recovery have both been increased.This paper discusses the changes made in the grinding circuit over the past few years and assesses their impact on operating costs and mill operations.IntroductionThe Williams Mine is located in the Hemlo area of northwestern Ontario, about 40 km east of Marathon. It is operated by the Williams Operation Corporation, which is equally owned by Teck Corp. and Homestake Canada Inc. Canada's largest gold mine, the Williams Mine commenced operation in October 1985 at a rate of 3, 000 t/d. Mill throughputs have typically exceeded design, averaging 3, 397 t/d in 1986, and 3,380 t/d in 1987. In August 1988, production was increased to 6,000 t/d. Throughput averaged 4,212 t/d in 1988. Production increased over the next few years averaging 5,477 t/d in 1989, 6,248 t/d in 1990, 5,987 t/d in 1991, and 6,283 t/d in 1992. With a recovery of over 95.5%, production for the first three quarters in 1993 is averaging 6,408 t/d.GeologyThe main ore body is a steeply dipping conformable zone of microcline-quartz-muscovite rock. It occupies the contact zone between hanging wall metasedimentary rocks and foot wall felsic quartz feldspar porpyritic rocks of possible volcanic or subvolcanic intrusive origin. The ore body is enveloped by muscovite schists which are thickest on the footwall side. The schist is enriched with tourmaline, pyrite, and green mica containing vanadium and barium. The ore body contains pyrite and barite which occurs mostly in laminae and bands to veinlets and breccia fillings. The gold occurs as free (native) grains, 20 to 30 µm in size at grain junctions of the microcline and quartz. Visible gold is rare in the main zone. There are commonly two or more discernable layers of economic value which are separated by weakly gold mineralized material which usually consists of muscovite schist, barite rich zones or sulphide poor-microcline-quartz rock and in the upper and western fringes by biotite rich schist with feldspathic fragmental characteristics. Molybdenite is a common constituent as very fine disseminated grains in the microcline-quartz rock and is most prominent with gold enrichment. Stibnite and realgar are two of the more common accessory sulphides."

Jan 1, 1994

By D. Rahal

All ores have an economic optimum particle size (Wills and Napier-Munn, 2006). This optimum value is often a compromise between the need to achieve adequate liberation for the efficient separation and recovery of valuable minerals and the energy consumed by the crushing and grinding process. It has been recognized for nearly three decades that grinding is the most expensive and least efficient stage in mineral processing (El-shall and Somasundaran, 1984). At that time it was estimated that grinding consumed between 25 and 70% of the total power required in mineral processing. This recognition drove the development of new stirred mill fine grinding technology in the 1990s (Jankovic et.al, 2003; Enderle et. al., 1997; Massey, 2008). These milling technologies have a reduced energy consumption compared to traditional fine grinding using ball mills. In the ?coarser? fine grinding applications this improvement is approximately 30%. The benefits increase to above 50% in ultrafine grinding applications (Lichter and Davey, 2006). Despite these gains, the US Department of Energy reported that an additional energy savings of up to 37% is possible in the minerals industry. They identified continuing research and development and implementing best practices in both grinding and materials handling as the primary avenues for the additional reduction (BCS, 2007).

Jan 1, 2011