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By C. R. van Staal, L. R. Fyffe, J. A. Winchester

"The geochemical characteristics of Early Paleozoic volcanic rocks are used to establish the following tectono-volcanic regimes in the northeastern mainland Appalachians: foreland continental rift volcanics; fore-arc ophiolites; island-arc volcanics; back-arc ophiolites; and back-arc ensialic volcanics. A Late Precambrian calc-alkaline magmatic arc bounds the Early Paleozoic volcanics to the southeast.The Precambrian calc-alkaline volcanics of the Caledonian Highlands, primitive Cambro-Ordovician oceanic-arc volcanics of the Eastern Townships, and within-plate Ordovician volcanics of the Miramichi Highlands can be related to a progressive change in mode of subduction from compressional Andean-type in the Late Proterozoic to extensional Western Pacific-type in the Early Paleozoic. The development of marginal oceanic basins in the Paleozoic created a more favourable environment for volcanogenic sulphide deposition, particularly in back-arc areas.The influence of tectonic setting is reflected in the distribution, size and metal ratios of the massive sulphide deposits associated with the volcanic rocks; deposits within intra-oceanic arcs in the Eastern Townships of Quebec average two million tonnes grading 2.2% Cu, 4.0% Zn, and 0.6% Pb, whereas those within a back-arc, ensialic rift near Bathurst, New Brunswick, average nine million tonnes grading 0.6% Cu, 5.4% Zn, and 2.2% Pb."

Jan 1, 1990

By H. Itou

A mechanism of scorodite formation at ambient pressure was investigated using transmission electron microscopy (TEM). Several types of iron arsenate were obtained from Fe/As solutions (Fe/As = 1.0 and 1.5) at pH values of 1.0, 1.25 and 1.5 at 95°C. Amorphous FeAsO4.xH2O was precipitated from solutions having a higher mole ratio of Fe/As at pH l.5, even after 9 hours. A few micron sized particles were observed by SEM, but these proved to be agglomerates of 10-50 nm amorphous FeAsarxH2O particles when observed by TEM. This means that supersaturation of the solution causes the acceleration of nucleation of iron arsenate hydrate. Crystalline scorodite was obtained from a solution in which the Fe/As mole ratio was unity, at pH l.0 and pH 1.5, after 3 hours. Almond-type crystalline scorodite particles were observed by TEM and most of them showed high crystallinity. The addition of scorodite seed material enhanced the crystallization. Amorphous FeAsO4.xH2O, however, was observed at the surface of the seeds during the early stages of precipitation. It is postulated that crystalline scorodite is not directly precipitated even on crystalline scorodite seeds, but requires a certain time to crystallize on the seed material.

Jan 1, 2006

By B. Kenney

For SOFCs, composite cathodes comprising an electronically conducting electro¬catalyst (e.g. LSM) and an conically conducting electrolyte material (e.g. YSZ) are considered to improve the electrochemical performance by extending the active reaction zone to regions beyond the electrolyte-electrode interface, as in the case of conventional cathodes. Interpretation of experimental data for composite cathodes, however, is complicated by the mixed conducting behaviour which results in local overpotential distributions throughout the composite cathode. These local overpotentials depend on many factors, such as the composition, particle size, porosity and thickness of the cathode. Reliable experimental data obtained from composite cathodes requires carefully crafted microstructural parameters in order to minimize the overpotential variations. A mathematical model of a composite cathode was developed using coupled differential equations describing transport and consumption of chemical species, oxygen ion and electrons. Simulations under various oxygen concentrations were used to obtain the reaction order of the oxygen reduction reaction. An apparent po2 dependency was then extracted from the predicted Rp vs P02 data. It was found that this p02 dependency can be significantly influenced by the micro-structural parameters of a composite cathode, primarily due to local overpotential distributions.

Jan 1, 2005

By S. D. Anitha Kumari, T. G. Sitharam

"The huge urbanisation of the world cities have invariably resulted in enormous requirements of development in transport and other facilities. As the surface space available for the construction of these infrastructural facilities is found to be inadequate, underground space is widely used recently for this purpose. However, to understand the behaviour of these underground structures through experiments under various loading conditions is not practically feasible. Hence numerical methods have widely been used to model these structures and predict their behaviour. The underground structures generally rest on soil or rock. Depending upon the properties of the material surrounding the tunnel, the deformation pattern and the stress contours vary. If the tunnels are placed on soft ground at a shallow depth, the presence of existing structures will add to the stresses and strains developed. In the case of densely populated cities, tunnelling operation may result in huge damages if the settlements are not within the desirable limits. This may in turn lead to the collapse of the tunnel along with adjoining structures. So the determination of stress levels associated with the yield of wall due to excavation is very important and necessary supports and reinforcement should be provided. Numerical simulations using discrete element method (DEM) and finite element method (FEM) have been done to understand the stress variation around single and twin tunnels. Studies are also done to quantify the improvement of stability on the introduction of lining around the tunnels in a weak cemented soil mass like sandstone.INTRODUCTIONMost of the life line facilities which include pipelines for electricity, water, communication, sewage, gas, roads, metro, utility tunnels etc in modern urban areas are underground structures. Infrastructure development is crucial for any city and sustainable development urges for underground constructions. Accordingly, there is already a great demand for tunnels to be constructed in difficult and crowded urban settings. This results in the fact that a large number of risk factors are associated with tunnels in these densely populated urban areas. These constraints pose threats to tunnel construction and also demands for minimal disturbance to the public and to the surrounding utilities. One of the most important factor to be considered while designing an underground structure is its stability with respect to the surrounding medium and adjoining existing structures. In this study, a comprehensive presentation of the numerical modeling of single and twin tunnels adopting two different methods is presented."

Jan 1, 2015

By Michel Mellinger

This paper presents preliminary results of a project aimed at obtaining a better understanding of the genesis of gold mineralization in northern Saskatchewan using lithogeochemistry. The properties sampled were: Weedy Lake (B Zone), Waddy Lake (Komis Zone) and Wedge Lake (Twin Zone) in the La Range belt, and the Rio mine in the Flin Flon area. Although the three vein-type deposits rudied here (Weedy Lake. Waddy Lake, Rio deposit) occur in different geological situation . a common gold deposition mechanism appears to have been active at all three location . Hydrothermal fluids carrying sulphur in solution in exces of gold thio-complexes moved through fractured host-rocks until they reached a site where pyrite precipitation wa triggered. As pyrite precipitated, the a/ aAu activity ratio of the fluid decreased. until a critical thre hold was reached and gold thio-complexe were de tabilized . resulting in the co-precipitation of gold and pyrite. The conditions triggering the precipitation of pyrite appear to have been: the abundance of iron in host-rock ~ilicates at Waddy Lake. and temperature decrease upon the reaching of structural trap~ at both Weedy Lake and the Rio deposit. with additional involvement of carbonatizationrelated reactions between fluid and host-rocks for the latter. In addition to pyrite precipitation , potassic alteration of the mafic metavolcanics is observed at Waddy Lake. while at the Rio deposit, the extensive carbonatization of the host-rocks i coupled with the deposition of chlorite that may have been remobilized from the altered mafic host rocks. In both the B Zone at Weedy Lake and the Komis Zone at Waddy Lake, a di tinct vanadium signature give some clue about the direction in which the hydrothermal fluids have moved . and the interpretation made is con istent with the geological ituation in either case. As for the Wedge Lake deposits, the distinct base metals-precious metals a sociation within a sulphide-facies iron formation confirm the volcanic exhalative origin of this gold mineralization .

Jan 1, 1986

By Doris Stüben, Geoffrey P. Glasby, Akira Usui, Jörg-Detlef Eckhardt, Bruce W. Mountain, Zsolt Berner

"Abstract - Twenty-seven marine Mn and Fe deposits chosen for their diverse origins have been analyzed for five platinum-group elements (PGE). The PGE display their highest concentrations in hydrogenous Mn crusts and nodules, lesser concentrations in diagenetic nodules and two to three orders of magnitude lower concentrations in hydrothermal Mn and Fe crusts. This difference of behavior reflects the well-known fractionation of Mn and Fe from the other transition elements, rare earth elements (REE) and PGE in hydrothermal Mn and Fe crusts. Pt, Ru, Rh, and Ir show a close geochemical coherence in these deposits and are differentiated from Pd which shows a negative anomaly with respect to the other PGE in hydrogenous and diagenetic Mn deposits. Pd is less depleted than the other PGE and the Pd anomaly much less pronounced in the hydrothermal Mn and Fe crusts relative to the hydrogenous and diagenetic Mn deposits.Stability field diagrams show no evidence for the oxidation of Pt to the tetravalent state in seawater as suggested by some authors. Rather, Pt appears to be present in seawater as Pt(OH)2 0. Adsorption of Pt(OH)2 0 onto the surface of Mn and Fe oxyhydroxide minerals may explain the enrichment of Pt in marine Mn deposits. However, Pt metal is the thermodynamically stable solid phase of Pt under seawater conditions and reduction of Pt(OH)2 0 to Pt metal, particularly in the vicinity of the oxygen minimum zone, is possible, although the high Ce/La ratios of the Co-rich Mn crusts do not support this hypothesis. Pt, Ru, Rh, and Ir display their maximum concentrations in the Corich Mn crusts. This may reflect, among other things, the low growth rates of these deposits. The stability field diagram of Pd shows that Pd is undersaturated in seawater and occurs in seawater dominantly as PdCl4 2- and anionic mixed-hydroxide-chloride complexes. This would explain the anomalous position of Pd among the PGE with respect to its uptake into marine Mn deposits. A study of inter-element associations indicates that the incorporation of the PGE from meteoritic dust plays only a minor role in the uptake of these elements into marine Mn deposits. The PGE remain the most poorly studied group of elements in marine Mn deposits. The data presented here emphasize the need for the establishment of a more comprehensive database of the PGE in marine Mn deposits in order to understand more fully their uptake into marine Mn deposits and their pathways through the oceans."

Jan 1, 1999

Abstract - A method to extract certain types of error variance information from paired data is presented, which constitutes an important tool for the calibration of Gy’s formula for variance predictions in the sampling of broken ores. The research summarized here complements publications by others on geochemical precision. The model and method proposed for treating duplicate data provide a practical means of deriving separate estimates of the purely analytical and aliquot sampling variances, from a unique set of duplicate assays, and, hence, proposes a new avenue for estimating effective mineral grain size.

Jan 1, 1998

By Michael McClintock, Bern Klein

"The high pressure grinding roll (HPGR) is a relatively new comminution technology that is gaining popularity for providing energy efficient solution to hard rock mining operations. A significant barrier to increased adoption of HPGRs is the present requirement for large amounts (excess of 2 tonnes) of sample for pilot testing, which may not be possible for most early stage projects. Three methodologies have been developed at NBK Institute of Mining Engineering at the University of British Columbia that may be used for sizing design optimization of HPGR circuits. All three methodologies use a MTS pistonpress to simulate HPGR comminution. These tests are capable of providing information regarding energy consumption, energy breakage relationship, reduction ratio, and product size distributions. In a period of slow economic development, these methodologies are critical to allow the industry to cost effectively understand the operational benefits of HPGR implementation. This paper examines previous test work done at UBC, as well as the development of an HPGR and piston press database for use in sizing and budgeting of HPGR circuits.INTRODUCTIONCurrently, the mining industry faces several challenges due to falling commodity prices and escalating costs. This has applied pressure for operations and projects to reduce costs and improve project economics. In addition engineering and metallurgical budgets have shrunk as companies try to conserve cash. Improvements and optimization of milling and comminution circuits is one area project economics may be significantly improved. By considering alternative comminution technologies such as the high pressure grinding roll (HPGR), energy operating costs may be greatly lowered. Many have begun to implement HPGR crushing as an alternative to semi-autogenous grinding (SAG) mills, including KGHM and, Sierra Gorda, and Freeport Mc-MoRan’s Cerro Verde operation. However, implementing new technology for companies with little or no HPGR experience may not be practical. An operator must consider both feasibility and economics of a circuit. Currently, consideration of the HPGR requires expensive pilot scale testing for which large sample sizes are needed. In addition, few laboratories are equipped for running pilot scale HPGR test programs. Furthermore, the results of pilot HPGR testing are dependent on the quality of the sample, which cannot represent variations across a deposit. In contrast, ball mills and SAG mills can be sized using the Bond Mill Work Indexes that require a small amount of sample which enable a number of different geological blocks to be tested to allow engineers to have a complete understanding of how the comminution circuit will perform for the life of mine."

Jan 1, 2016

By J. Weyer

"Threshold limit values (TLVs) for NO and NO2 have been reduced drastically in 2016. Emissions of diesel engines have been reduced in the last decades. In many cases this is not enough to stay under required TLVs. Remaining gas in the blasted rock is set free during loading and transport. At the end, only a combination of measures with a further reduction of diesel motor emission, replacement of diesel driven vehicles by electric equipment, new explosives with reduced NO and NO2 production, changes in the organization and effective ventilation measures can ensure that the TLVs are not exceeded. KEYWORDS Blasting fumes, Gas measurement, Gases, Threshold Limit Value (TLV) SHORT HISTORY OF EXTRACTION OF RAW MATERIALS The extraction of raw materials—besides haulage and transport—is one of the main processes in mining and tunneling. The principal objective is the creation of a material which can be transported from the place of extraction (face) to another place for further processing or direct use. Furthermore, recently a desired grain size distribution has become more important. A revolutionary development took place in the course of centuries. With the use of primitive tools like shovels, chisel and iron in the 12th/13th century in Freiberg, ore mines performances of 12 16 cm per man and week could be reached in cross section of 1 m2. Recalculated with an assumed average density, this is a performance of 0.006 t/h. With the introduction of black powder into the subsurface in 1627 by Caspar Weindl in the city of Schemnitz (nowadays Banska Stiavniza), performances increased. A part of the hard extraction work was now done by the explosives. But first in 1767, it was recognized in a mine in Freiberg that more than 1 hole could be drilled (hand-drilling) and blasted. This discovery was so important that a plate with an inscription describes this event. The performance was increased more than 10 times compared to the extraction with chisel and iron. Since this time, drilling and blasting has been the most important innovation force for the increase of the performance. Problems with noxious blasting fumes led to the construction of the first tunneling machine by Henry Joseph Mouse in 1846 (beside shield machines from 1825) (Stack, 1982)). The triumph of mining machines started only interrupted by the two world wars. The development of hard rock boring machines (including roadheaders) was enforced after 1945. An almost complete extraction by hard rock boring machines was predicted for the turn of the millennium. And, indeed, one could find powerful machines in Canadian potash mines which were able to mine 1000 t of potash in 1 hour. That was an increase in the performance by several powers of ten. But already in the 1980s, it was recognized that the ambiguous goal could not be reached. Problems with wear of the tools, radius of curves, minimal length of openings, size and costs of the machines led to a more or less parity between drilling and blasting on one side and machines on the other side, of course depending on the field of use. For this reason, drilling and blasting is still the dominant extraction technology for raw materials. In the development of drilling and blasting technologies, the focus was mainly on the improvement of the safety but also on the increase of the blasting power and performance. Examples are tests for the increase of the power of ANFO explosives which were carried out in the 1970s and 1980s, in the former teaching mine “Reiche Zeche”. In the 1980s vehicles for loading and mixing of ANFO explosives have been developed, the blasting cab sensitivity of ANFO has been increased, emulsion explosives have been developed and up to now further developments like PMMA LOX EX explosives are on the way. At the turn of the millennium, 6 million t/a explosives have been used worldwide. Annual rates of increase of 2% have been predicted (Weyer, 1999). Meanwhile the focus changed from an increase of the performance"

Jan 1, 2019

By G. E. Murray

Introduction The development of the method for recovering zinc from lead blast-furnace slags at Trail is interesting and important in the history of both lead and zinc metallurgy. Many efforts have been made in the past, particularly in Germany, to make economic recovery of zinc from slag. The Pape process on Oker slags was successful and is being worked at the present time. More recently the Waelz process has come to the fore. One effort at Cockle Creek, Australia is of special interest in view of subsequent developments. This was re-ported in detail by Guy Courtney in 1920(1). It deserves mention as it shows how, as in many other instances in the metallurgy of lead and zinc, important pioneering work has been done in Australia. These experiments, started over 25 years ago, comprised reduction with coke, powdered coal, and eletrothermically in carbon-lined furnaces. The so-called slag eye furnace, which volatilised the metals from a blast furnace running with a hot top, was also investigated.

Jan 1, 1933

By M. Drozd, M. Somppi, Z. F. Yamak

"Activated carbon is used to recover gold in solution from various gold circuit designs including carbon-in-pulp (CIP), carbon-in-column (CIC), and carbon-in-leach (CIL) processes; however, the characteristics and durability of the activated carbon can vary. Handling this medium poses challenges when fine carbon losses occur in the CIP, CIC, and CIL circuit discharge. Additionally, adsorbed gold on carbon fines can be lost to the discharge. The studies in this paper aim to standardize laboratory methods for testing activated carbon qualities specific to the gold industry. The investigation focused on size distribution, platelet content, wet attrition resistance, gold kinetics, and gold loading. The results demonstrate that a 96 h attrition test and a gold loading test form a basis for ranking carbon samples based on expected economic performance of the plant. ¦ KEYWORDS Activated carbon, Carbon attrition, Carbon ball-pan hardness, Carbon evaluation, Carbon fines loss, Carbon turnover RÉSUMÉ Le charbon actif sert à récupérer l’or en solution dans diverses conceptions de circuits de broyage d’or, notamment les procédés de charbon en pulpe (CIP, de l’anglais carbon-in-pulp), de charbon en colonnes (CIC, de l’anglais carbon-in-column) et de lixiviation au carbone, ou charbon au lixiviat (CIL, de l’anglais carbon-in-leach). Toutefois, les caractéristiques et la durabilité du charbon actif varient. La manipulation de ce milieu pose des problèmes dès lors que se produisent des pertes de carbone à grain fin dans la décharge des circuits CIP, CIC et CIL. Par ailleurs, l’or adsorbé sur les fines de charbon peut se perdre durant la décharge. Les études réalisées dans le cadre de cet article visent à normaliser les méthodes de laboratoire pour tester les qualités du charbon actif spécifiques au secteur de l’exploitation aurifère. L’étude portait sur la composition granulométrique, la teneur des plaquettes, la résistance au frottement humide, la cinétique de l’or et l’adsorption de l’or, à savoir la capacité du charbon actif à adsorber l’or sur une période spécifique. Les résultats ont montré qu’un test de frottement de 96 heures et un test d’adsorption de l’or constituaient le fondement d’un classement des échantillons de carbone du point de vue économique en vue de trouver l’échantillon générant le meilleur rendement des usines. ¦ MOTS CLÉS charbon actif, dureté de la batée des billes de carbone, évaluation du carbone, perte de fines de charbon, renouvellement du carbone, résistance au carbone INTRODUCTION The use of activated carbon for gold extraction is among the most common processes in modern gold-milling plants. Although gold losses are inevitable in the milling process, operators often focus on reducing gold loss to tailings. One focus area that is important in gold recovery is optimizing the use of activated carbon and examining potential changes to realize improved economics for carbon-in-pulp (CIP), carbon-in-column (CIC), and carbon-in-leach (CIL) processes. Granular activated carbon (Figure 1) is widely used for gold extraction from slurries containing gold-cyanide complexes. This activated carbon—produced from charred coconut shells—is well suited because of its adsorption kinetics, gold-loading capacity, attrition resistance, and ability to be reactivated. The extraction process using activated carbon technology involves adsorbing gold from an aqueous leachate; the gold is dissolved using a cyanide salt and consequently a dicyanoaurate(I) ion (Au(CN)2–) is formed, as noted in equation 1."

Jan 1, 2019

By T. Yalcin

"This paper describes a study carried out to determine the behaviour of Cu-Ni mill tailings during natural and flocculated sedimentations. Settling rates were examined under various conditions through standard laboratory batch settling tests. Flocculation was found to improve the sedimentation process significantly, but its effectiveness depended on several factors such as the pulp density of feed, the type and dosage of flocculating reagent, the pH of pulp, the intensity of mixing flocculant with tailings sand the use of stirring during compaction.It is shown that. for a sedimentation process taking place under given conditions, underflow pulp density (P) and sedimentation time (t) are related by the expression P = k.tn, where k and n are constants. A procedure is described for the use of this expression in thickener size estimations. IntroductionOften a considerable proportion of the mill feed is disposed as tailing, because the valuable mineral recovered usually forms a small fraction of the ore. Moreover, wet concentration processes result in liquid wastes that are several times more voluminous than the ore itself. Thus the handling and disposal of these wastes become an important cost item in most milling operations. The principal unit process employed in tailing disposal is solid-liquid separation which is traditionally carried out by gravity sedimentation in thickeners or tailing dams. Direct disposal of tailings into tailing dams is common practice where sufficiently large areas of land are available but it frequently becomes more economical or practical to densify the tailings in a thickener beforehand.Successful thickener operation depends essentially on correct thickener size which is in turn governed by the sedimentation characteristics of the tailings handled. Flocculation can aid the sedimentation process and bring about significant savings in thickener size, but its effective use requires a detailed study to identify the optimum conditions of application."

Jan 1, 1988

By M. A. Cristovici

"In the last few years CANMET has taken the initiative by investigating the possibility of reactivating abandoned old tailings dumps (for gold production) which would be of potential economic interest across Canada.Nova Scotia was one location where numerous gold processing plants were in operation in the early days. A survey of the old tailing dumps was conducted and some areas of interest were selected. Among them were the sites identified as Lake Charlotte and Forest Hill.Ore samples collected from those tailing ponds were examined for their chemico-mineralogical composition and investigated for gold extraction. Three technological concepts were tested: flotation, direct cyanidation and flotation followed by cyanidation of the flotation concentrate. The combined flotation- cyanidation variant, yielding a gold recovery in the range of 90%, was recommended for processing the tailing ore.Pilot-plant runs carried out by a commercial testing firm (data not included in this paper) confirmed the conditions and results of CANMET's technology.As a result a consulting-engineering firm designed the processing plant which is expected to begin operation in the near future.IntroductionIn the last few years CANMET initiated a research program to locate in Canada, abandoned gold-silver tailing dumps of potential economic interest and to explore the possibility of recovering the precious metals from these old deposits by modern techniques.A literature search revealed that the first production of gold in Canada was recorded in British Columbia in 1858 and that, production prior to 1900 was essentially restricted to the Yukon, British Columbia Nova Scotia and Quebec. During the early 1900s, Ontario, Manitoba and the North West Territories joined the ranks of the gold producers.Early milling and recovery methods were mainly restricted to sluice boxes, gravity stamps and amalgamation. Froth flotation and cyanidation were introduced at about the turn of the century.Recovery of gold at plants using flotation and cyanidation was very good, therefore the tailing ponds resulting from those operations would, very likely be of little or no economic importance. However, the gold recovery at the amalgamation and gravity concentration plants appeared to be lower, indicating the possibility that significant amounts of recoverable gold would have been carried with the tailings to the dump sites."

Jan 1, 1986

By M. J. Bunyard, I. Mullany

"The pressures on mine operators are varied and constantly increasing, declining ore grades, more remote locations and increasing environmental concerns. The challenge facing the mining industry is to find new ways to process ores which are more efficient and less costly and to control them better. This requires a fundamental rethink in the way in which projects are developed, it is no longer good enough to just copy what has been done before. The process starts with the surfacing of the primary underlying constraints for the project and the establishment of new design principles to address these. The generation of new design criteria and processing options follows the well-established path of brainstorm, focus, brainstorm, refocus and economic evaluation. The aim is process intensification, doing more with less, as early as possible and better use of plant data. The methodology requires a complete geometallurgical understanding of the orebody to facilitate the evaluation of viable processing options. Examples of early stage processing considered are blast optimization and ore sorting to reject waste and reduce processing costs. Mine to mill integration, which starts with the optimization of blasting for separation, offers another way to significantly reduce operating costs and increase efficiencies.VALUE MODEL The mining industry has, over the past decade, experienced an unprecedented cycle of increased commodity prices and investments in new mining and processing capacity followed by a market collapse and fall in prices and confidence. Whilst there are now some signs of a return to investment in new capacity, the metals and mining industry is still wrestling with eroding margins and investor demand that companies achieve the full value of their assets. Metals and mining companies must work on all available levers to deflate their costs back to more manageable levels and capture the full value for their products. Levers include optimizing external spending, re-thinking make vs buy trade-offs, debottlenecking, improving availability and operating time of key equipment, more accurate and agile planning processes, value-in-use optimization, and tighter performance management. Figure 1 is a diagrammatic representation of a company’s value model. The three goals of a successful mining enterprise are social acceptance, environmental acceptance and sound financial returns to shareholders. Some of the KPI’s used to measure the success or not of achieving these goals are shown in the translation layer."

Jan 1, 2017

By M. Olper

Commercial copper plants utilising current best technology are tending towards ever increasing scale, the concentrate requirements of which cannot be serviced by a single mine. In such cases, raw materials are often sourced from different mines, which are often far away from the production plant and usually with very different ore compositions. These plants tend to use the traditional thermal processing steps of smelting, converting to blister, anode refining and casting to produce copper anodes that are then refined in a sulphate electrorefining operation. These thermal steps are in some cases troublesome and require the recycling of considerable quantities of material. A new process was developed taking advantage of the latest developments of the first thermal step of the classic loop, the smelting of copper concentrate to high grade matte or white metal using the well known AUSMELT Top Submerged Lancing Technology. The white metal and/or high grade matte product is converted directly to copper cathode and elemental sulphur using the ECUPREX technology. This approach provides an alternative to the traditional thermal steps of conversion to blister, thermal refining and anode casting, which, especially at small scale, can be expensive and problematic in terms of technical and environmental concerns. At first evaluation, this new process can be operated at low capacities, thereby, changing the traditional approach to copper production. Instead of taking the concentrate to the copper production plant, the copper production plant is taken close to the mine with some significant economic and technical advantages.

Jan 1, 2007

By D. Berkley

The competitive pressure, the increasing energy costs and the environmental regulations impose substantial challenges to the definition of power plants serving industrial facilities. Primary mining and metallurgical facilities have even greater challenges because they are often located remotely, where the fuel supply can be a problem. The absence of a power grid imposes challenges on the reliability and quality of the electrical energy supply. In addition, the facilities usually require thermal energy too, not only electricity. This paper outlines these types of issues that have to be considered when defining a cogeneration plant integrated with a remote located industrial facility. It is addressed to the ?users? of cogeneration facilities rather than to ?generators.? The focus is on analyzing various site-specific design considerations to pre-screen the feasible cogeneration alternatives and minimize the number of options to be analyzed from the economics viewpoint. This approach could substantially reduce the engineering effort and schedule during the feasibility phase of the project. Part 1 of this paper (Berkley, 2003) provides a basic understanding of cogeneration plants of various types to establish a common language. Part 2 outlines and analyzes specific issues that could be of help in the definition of cogeneration installations for remotely located industrial facilities, with a focus on mining and primary metallurgical plants.

Mar 1, 2003

By M. Razavi

This investigation is part an overall investigation on the effect of sodium silicate on the properties of the mine backfill. The results presented will highlight the effects of binder dosage, sodium silicate concentration, and pulp density on the performance of stabilized sodium silicate-fortified sand pastefill samples. Binders implemented for these experiments consisted of cement, slag, and sodium silicate in various proportions. The binder content varied from 3 to 9 wt% (total dry weight) and sodium silicate concentrations was adjusted at 2 and 4% (total binder weight). Moreover, pulp density was adjusted to values between 80 and 83 wt%. For completion of the aforementioned phase of the study, series of sand pastefill test samples were prepared and cured for 28, 56, or 120 days, in a humidity and temperature controlled room. The specimens were tested for uniaxial compressive strength (UCS). In order to study the effect of sodium silicate on microstructures of stabilized sand pastefill, some MIP tests were conducted on selected samples. The results of this investigation evaluate the use of sodium silicate as a partial replacement of cement in stabilized backfill. The analyses of the results demonstrates that, with an increase in binder content and a decrease in pulp density, the UCS values of sand pastefill samples increase. Further more, the results indicate that increase in addition of sodium silicate does not necessarily result in higher strength minefill.

May 1, 2007

By W. A. Gow, V. F. Harrison, H. H. Mccreedy

A semi-continuous bench-scale, countercurrent-decantation system, in which the ore was moved countercurrently to the flow of leaching solution, was used to conduct acid-leaching tests on uranium ore with the aid of bacteria. This method, which involved a low degree of agitation, was successfully applied to extract uranium in practical amounts and at practical rates. The acidity of the leaching solution was the main variable investigated in the bench-scale tests done at the Mines Branch, Ottawa. Over 90 11er cent of the uranium was extracted in a six-stage system. The results suggest that this technique could be ap-plied in the utilization of bacterial oxidation and leaching as a method for treating Elliot Lake uranium ore.

Jan 1, 1969

By G. K. Livingstone

"This paper describes mining conditions in the Kaiser surface and underground coal mines in the Sparwood area of southeastern British Columbia. Mining methods are explained, major mining equipment is listed and production rates are given. Reclamation methods and results are described. Data are given on safety organization, safety practices, accident statistics, safety inspections and absenteeism. IntroductionThe coal mines of Kaiser Resources Ltd. are located in the southeast corner of British Columbia. All present mining is in the Sparwood area, where operations began in 1900 with the opening of the Michel underground mines. Production increased rapidly at first, then went through successive periods of depression and prosperity. Following are the highlights. Quantities are in clean, short tons. 1900 1910-Rapid rise in production to a peak of 512,000 tons in 1910.1911-1924-Reduced rate of production, averaging 238,000 tons per year, due to strikes, manpower shortage during World War 1 and post-war depression.1925-1929-Second prosperous era, with production climbing to 383,000 in 1929.1930-1933-Second depression-production falling to 231,000 in 1933.1934-1942-Third era of prosperity-production recovered and new production peak of 908,000 tons reached in 1942 during World War II.1943-1956-Corttinued prosperity, with production averaging 720,000 tons per year.1957-1959-Loss of markets to oil and gas; production dropping to 517,000 tons in 1959.1959-1969-Recovery due to new coal market in Japan; production climbing back to the 900,000-ton level.197D-Present-Jump in production reaching 6.6 million tons in 1975. This was the result of the purchase of the property by Kaiser Steel, who negotiated a large-volume, long-term contract with Japan, built the coal port of Westshore, reduced shipping costs by instituting unit trains, and reduced production costs by investing in large-scale mining machinery and a new preparation plant."

Jan 1, 1980

By A. Mortazavi, H. Nasab

"The storage of petroleum products on surface is associated with difficulties and limitations. Accordingly, a viable alternative is to excavate underground space in rock and provide a safe way for oil storage. Soft rocks such as salt domes provide good conditions from environmental and operational points of view. The potential for high-volume storage and low permeability are encouraging aspects for oil storage in salt domes. The complicated shape of oil storage caverns, the complicated behavior of salt rocks, and boundary conditions associated with large underground openings are important challenges in the design of oil storage caverns. The aim of this paper is to delve into the stability and deformation mechanisms of storage caverns created in salt domes. A parametric study of effect of the cavern size, orientation, and depth is carried out as a function of ground stress regime. The stress field and deformation mechanisms are calculated for each case aiming at providing design guideline.INTRODUCTIONOne of the most popular soft rock that is used for underground oil storage is salt rock (halite). Salt rock has particular characteristics which makes it suitable for oil storage. Salt rock has several advantages for oil storage, these are: high production rate, high degree of availability, short filling period, high level of safety and low permeability [1]. The low permeability is due to the fact that halite deforms in a plastic manner over a wide range of pressures, temperatures and rates of deformations. This property prevents joints or gaps opening up in the formation, thus preventing the escape of liquids and gases. The state of stresses around the cavern in salt domes depends on the depth of the cavern (initial stresses and rock mass temperatures are increasing with depth), the in-situ stress state within the rock mass, and the internal oil pressure in the cavern. Furthermore, the salt creep behavior and the cavern geometry are influencing stress redistributions during loading and unloading phases.Rock mass structure, cavern geometry, material behavior and cavern operation patterns require complex physical models to describe the load-bearing behavior of the geomechanical structure. The determination of stress, strain and deformation in the rock mass due to cavern construction and operation are complex problems which can be evaluated employing numerical tools such as the Finite Element Method (FEM) or Finite Difference Method (FDM). In this study, the finite difference method was used to evaluate the behavior of oil storage cavern in salt rock."

Jan 1, 2015