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By D. R. Nagaraj, Napoleon Tercero, Raymond Farinato

"Dialkyl dithiophosphinates are a unique class of collectors used in the flotation processing of Cu- Au, complex polymetallic, and precious metal ores. They are known to exhibit higher selectivity and flotation rates for Cu sulfides, galena, and precious metal values in the presence of other gangue sulfides, than the traditionally used collectors such as dialkyl dithiophosphates and xanthates. They also have an attractive HSE profile. Historically, for most reagents used in flotation, practical methods of chemical usage have preceded detailed scientific understanding of how they work; this applies also to dithiophosphinates. Recently, however, the research community has shown significant interest in this chemistry, and numerous studies have been conducted to understand the fundamentals of interactions of dithiophosphinates with several base metal sulfides and precious metals. A critical overview of fundamental studies, laboratory ore flotation, and plant practice of dithiophosphinates is given in this paper. The findings from these studies are rationalized by drawing on concepts and insights from coordination chemistry. A comparison between the interactions of dithiophosphinates and dithiophosphate analogues with various minerals and metals is also made throughout. We believe that such a critical analysis/overview will be invaluable in identifying gaps in knowledge regarding this important set of compounds, which in turn may provide a basis for the design of studies that can advance understanding of their flotation performance. INTRODUCTIONDialkyl dithiophosphinates are an important member of the class of thiophosphorus compounds that are widely used in flotation beneficiation of base metal sulfide and precious metals ores. Ore flotation studies and plant practice experience have demonstrated their unique properties compared to the traditionally used sulfide collectors such as dialkyl dithiophosphates (their closest analogues) and xanthates. The predominant dialkyl dithiophosphinate used in the industry today is sodium diisobutyl dithiophosphinate (a.k.a. DIBDTPi; AEROPHINE® 3418A1 introduced by American Cyanamid in the late 1970s; Figure 1). The widest application of AEROPHINE 3418A is found in complex polymetallic ores of Cu-Pb-Zn-Ag, Pb-Zn-Ag, and Cu-Zn where it has demonstrated improved recoveries (and flotation rates) of sulfide minerals of Cu, Pb and Ag and selectivity against sphalerite, iron sulfides, and penalty elements (Mingione, 1991). More recently, it has also been used as a collector in the processing of Cu, Cu-Ni, and Ni ores (Guitard et al., 2015) and Cu-Au ores (Dunne, 2005). In general, it is considered to possess the collecting power of xanthates together with the selectivity (against pyrite) of dithiophosphates (Hope et al., 2003a, 2003b), while requiring lower dosages than those reagents (Guitard et al., 2015). AEROPHINE® 3418A has also found an important place in several targeted formulations to improve flotation performance. Other benefits of this reagent include higher flotation rates, having a better HSE (health, safety and environmental) profile2 compared to xanthate, and exhibiting resistance to hydrolysis and oxidation (Guitard et al., 2015). Compared to diisobutyl dithiophosphate (DIBDTP), which is known to contribute significantly to frothing (often leading to over-frothing), AEROPHINE® 3418A is observed to be froth neutral (i.e. it does not alter the existing froth characteristics). As is the case for other flotation reagents, the design and application of AEROPHINE® 3418A has preceded understanding of its behavior at a fundamental level. Recently, there have been numerous fundamental and application studies designed to develop a better understanding of its interaction with sulfide minerals and precious metals. This should help in optimizing its application in flotation and provide a basis to better rationalize its unique selectivity and activity characteristics. Past studies on the interacti"

Jan 1, 2016

By H. W. McGerrigle

The St. Lawrence Valley of Quebec lies in an earthquake danger zone, the main centers of which appear to be Murray Bay, Montreal and Quebec City. This paper compares conditions in this "earthquake zone" with those of other, currently more active, earthquake zones of the Earth, deals briefly with the mechanics of earth movements and outlines the means by which earthquakes are re-corded. An account is then given of the earthquakes which have occurred in the Province of Quebec -from that of 1534-35 to the earthquake of 1925. A 50-year frequency cycle is implied. The general conclusion is that the 11robability of future earthquakes should be given more consideration in the construction of buildings in the St. Lawrence Valley region.

Jan 1, 1966

By Raymond S. Suglo

Increasing depths of mining coupled with the mechanization of most underground mining operations have led to growing complexities in the mine environments with the production of large amounts of gaseous and particulate contaminants. Over 100,000 lives have been lost due methane gas and dust explosions in coal mine workings in both the US and Canada since 1900 when the methane concentrations exceeded stipulated levels in some workings. There is therefore the need to constantly assess and evaluate the performance of the existing mine ventilation systems to maintain safe and acceptable mine environmental conditions in workings to meet the requirements of federal, state or provincial laws. This paper uses the results of continuous monitoring of methane concentrations conducted in selected development headings in coal mines in North America to assess the effectiveness of existing auxiliary ventilation systems in the control of methane gas concentrations in the headings. The results show that the average quantities of fresh air required to dilute, disperse and remove methane concentrations within set levels of one minute varied from 5.43 m3/s in Heading #11 (Condition 2) to 27.97 m3/s in Heading #10 (Condition 1) and this represented about 12 to 26% of the quantities required to dilute methane concentrations to safe levels within one minute under Condition 1 and 32 to 139% under Condition 2. The average dilution times in the headings studied were less than 8 minutes. The forcing efficiencies of the auxiliary systems in the headings studied ranged from poor to excellent implying that the auxiliary ventilation systems were capable of controlling the methane concentrations below statutory levels but may not be able to cope with large and unusual methane concentrations in the headings. This study is important to prevent methane gas and coal dust explosions in coal mines and thus avoid any loss of lives and property.

May 1, 2001

By John F. Kearney

After World War II Canada?s fledgling iron ore mining industry developed along the ill-defined Quebec / Labrador border near what became the town of Knob Lake and later renamed Schefferville. For about twenty-five years from 1954 to 1982 Schefferville was the epicenter of Canadian iron ore mining and the foundation of the Iron Ore Company of Canada. In about 1982 it all stopped, shut down, the miners moved away leaving behind a mining camp and a ghost town that have sat idle and almost deserted for the past twenty five years. The story looks set to change with new plans to re-establish the former mining operations in the historic mining camp. This paper outlines the proposal to reactivate the direct shipping iron ore operations in the Schefferville mining camp, and specifically the properties in Labrador owned or optioned by Labrador Iron Mines Limited (?LIM?), a wholly owned subsidiary of London Stock Exchange listed Anglesey Mining plc. The Schefferville mining camp is located about 1,000 km northeast of Montreal. There are no roads connecting the area to southern Labrador or to Quebec. Access to the area is by rail from Sept Iles to Schefferville, or by air from Montreal and Sept Iles.

May 1, 2007

By Engineering Staff

"This paper briefly describes the surface mining operations at Smoky River Coal Limited (formerly McIntyre Mines Limited, Coal Division) near Grande Cache, Alberta. Geology, mining methods, equipment, manpower and reclamation and environment topics are discussed. Introduction Smoky River Coal Limited, a wholly-owned subsidiary of McIntyre Mines Limited of Calgary, Alberta, produces a high quality, low sulphur, low volatile bituminous coking coal from its minesite located at west central Alberta, approximately 20 km from the town of Grande Cache and 360 km west of Edmonton (Fig. I). Most of the coal property is contained within a 30 km long by 20 km wide block lying in a northwest-southeast direction in mountainous terrain. Elevations vary from 920 m above sea level ·at the plant·site on the Smoky River to 1,600 mat the upper levels of current and proposed mining operations. A total of 11 coal seams have been identified. Only two seams, No. 4 Seam and No. 10 Seam are currently considered of major commercial value. These seams are recovered by both surface and underground mining. Approximately 650Jo of the plant feed coal is obtained from a conventional shovel/truck surface mine; the balance is mined underground by room-andpillar methods. Background In 1921, McIntyre Mines Limited acquired some 26,300 hectares of coal leases at Smoky River. Since then the leases were explored intermittently by different companies including McIntyre. In the mid-l 960s exploration activities increased and several feasibility studies were completed. Development of the area was undertaken by McIntyre in 1968 and, in August of the same year, a long-term contract was signed with a consortium of Japanese steel mills for delivery of 2 x 106 t of clean coking coal for 15 years commencing in 1970. Production was focussed on underground mining, using longwall methods."

Jan 1, 1985

By Zhao-shu Sun, Mao-lin Li, Rui Cui

"Fankou Mine produced pyrite concentrate (PC) from the tailings by rougher and scavenger flotation before 2010, with the averaged S grade of 40.54% at recovery of 79.34%. The economic value of PC increases with its purity. When the S grade of PC feed to sulfuric acid plant is over 47%, the pyrite cinder can reach a high Fe grade of 62% which can be directly sold as iron concentrate. On this background, engineers in Fankou Mine set the PC with S grade of 47% as target.The direct way to reach that goal was to clean the rough PC Lots of cleaner flotation tests were conducted and the balanced strategy was to produce two ranks PC with the S grade of 47% and 38% at the recovery of 44.20% and 36.17%, respectively. The dual-product flowsheet was successfully industrialized in 2010. In order to further improve the quality of PC, gravity separation was introduced into the flowsheet. The numerical simulation and tests were performed to evaluate the beneficiation and classification performance of hydrocyclone as it treated the rough PC. It was shown that the underflow of hydrocyclone could addressed the S grade of 47%, and the finer size overflow could also be enriched to PC with S grade of 47% by flotation, due to the contained pyrite with well liberation. That is to say, with the ingenious addition of hydrocyclone to treat the rough PC, single PC with S grade of 47% could be achieved. The new flotation-gravity separation flowsheet was commissioned to produce the high quality PC at the end of 2013. The averaged S grade is 47.20% at recovery of 82.26% in 2014. The two steps flowsheet improvement maximizes the enterprise benefits and leads to simple utilization of Fe in pyrite in downstream industry."

Jan 1, 2016

By V. Silins, J. Tameling, R. C. Cook

"A general review of the wear problems encountered in the mining industry is presented and current practice for wear protection systems is reviewed.A new concept in wear protection is described and compared to alternative systems. Utilization of conventional stud welding technology and abrasion-resistant wear studs is shown to be a useful supplement or substitution for conventional techniques. Operating life of many consumable components can be increased by up to four fold. The technique is highly portable for on-site field application or can be used in shop environments.Details of the wear protection model and metallurgical properties of the wear component are also reviewed. Laboratory wear test data are reported to supplement field test performance or to screen potential alloy compositions for application in specific environments.A Review of Wear Problems and Wear ProtectionThe emphasis on wear protection has increased in recent years because of the increasing impact of the potential savings in time, materials and energy. A 1974 German study concluded that attention to tribology, the study of friction and wear, could save approximately $3 billion, and a study in the U.S.A. at the same time put the figure at $100 billion of which $20 billion was for materials. A more recent American study put the energy component of the potential savings at $16.2 billion per year. A recent British study estimated the total potential savings in the U.K. as $1 billion per year, with the largest savings coming from lower maintenance and replacement costs ($500 million), lower breakdown costs ($250 million) and increased machinery life ($200 million). These three major contributions are all potentially accessible in the mining industry."

Jan 1, 1982

"THE GOLD MINING REGION of Timmins - Val d'Or, astride the Ontario-Quebec border, has long suffered the effects of rising costs during the period when the U.S. price of gold was fixed at $35 per ounce. The federal government, through such assistance programs as that set up under the Emergency Gold Mining Assistance Act (E.G.M.A. ) in 1948, has subsidized the gold mining communities of Canada for a total of over $300 million to March 31, 1972. Of this, the Timmins - Val d'Or ""gold belt"" has received more than twothirds, almost $220 million. In spite of these continuous subsidies, only 12 gold mines are still producing in the Timmins - Val d'Or belt, whereas 71 mines have received subsidies for some period since 1948. The vital question is how the communities previously supported principally by the mining of gold will survive in the long run, as many of the older mines either are already depleted or are quickly becoming so. With the rising price of gold on the 'free market in recent months, prospects look a little brighter for the time being. Even with the increased official price of U.S. $38, which made the maximum government subsidized price about $48 per ounce, gold mines have not sold any gold to the government since January 1972 and some stopped doing so months before that. The populace of the region has not been enthusiastic about relocation nor about extensive retraining programs in the past and probably will not be in the future. So, when the remaining gold mines become depleted, the old problems of community depression will remain."

Jan 1, 1973

By Fazel G. Jahromi, Ahmad Ghahreman

"Atmospheric leaching of arsenical sulfide concentrates, and especially enargite leaching, is one of the major challenges within the current hydrometallurgy research area. One of the recent processes developed for sulfide ore treatment is atmospheric leaching in the presence of AF 5, a novel carbon based catalyst, in chloride media. The quantification of different ion concentrations during enargite and chalcopyrite leaching in the presence of AF 5 were measured and the copper recovery rates of 97% and 96% were achieved. The objective of this study is to shed light on the leaching mechanism of the sulfide minerals. Based on the speciation of iron and copper by different titration methods, the AF 5 catalyst first reduced ferric and cupric ions and then oxidized ferrous, cuprous and arsenite ions. The oxidation behavior of the catalyst enhanced the leaching results. The AF 5 catalyst also caused in situ precipitation of scorodite in leaching experiment. The speciation of ions in the chloride media were simulated by Phreeqc software and the dominant species in the leaching solution were determined.INTRODUCTION The hydrometallurgy of copper has been used for several decades and is still under development for more complex copper sulfide minerals due to their refractory nature. Most of the world copper reserves are in the form of chalcopyrite; and enargite is known as a major contaminant. Pyrometallurgical processing of enargite is problematic due to arsenic release and tight environmental regulations (Peacey, Gupta, and Ford n.d.). On the other hand, lower recoveries in the hydrometallurgical processes make most of these processes impractical. Several leaching processes have been developed for the leaching of copper sulfide minerals. Some of the common treatment methods for copper sulfide minerals are pressure oxidation leaching (POX), CESL, Activox, Albion process, atmospheric ferric sulfate leaching, Dynatec, Galvanox, and Hydro copper process (Barr, Defreyne, and Mayhew 2005; Conner and Anderson 2013; D.G. Dixon, Mayne, and Baxter 2008; Hyvärinen and Hämäläinen 2005; Palmer and Johnson 2005; Rohner, Bartsch, and Ngoviky 2012). Generally it is not viable to dissolve copper sulfide minerals at ambient pressure in mild acidic media and dissolution of these refractory minerals requires harsh conditions in the presence of oxidant materials (Biswas and Davenport 2002). Ferric ions have been used as an oxidant in different media in copper sulfide leaching processes. U. S. Bureau of Mines’ Reno Metallurgy Research Center in 1969 (Haver and Wong 1971) investigated ferric chloride leaching of chalcopyrite. Dutrizac and Macdonald (1972) studied ferric sulfate leaching of enargite. Based on this study, 50% of the copper was recovered at 80-85 ºC after 7 days of leaching. Cupric ions can also be used as an oxidant. Parker, Paul, and Power (1981) stated that cupric/cuprous couples are more effective than ferric/ferrous couples in chloride media because they are rapid and reversible. The Hydro Copper process leaches copper sulfide minerals in chloride solution catalyzed by cupric ions. In this process, copper dissolves as cuprous ions in presence of cupric ions (about 10 g/L) (Hyvärinen and Hämäläinen 2005)."

Jan 1, 2017

By R. H. Chambers, L. Alderman

"The paper is introduced by an examination of the factors which have led to the present world-wide interest in charging preheated coal to coke ovens. Various possible methods of preheating and charging the coal are considered and reasons are given for the adoption of the methods chosen. The design of the pilot preheating plant for the British Coke Research Association is discussed and its influence on the design of the commercial units for the South African Iron and Steel Corporation's plant at Pretoria and the British Steel Corporation's plant at Beighton are explained. Operating procedures on the pilot plant are described, particularly with regard to the charging of the preheated coal. Design features of the commercial installations, engineered by Simon-Carves Ltd., are explained with emphasis on the built-in safety of the control system of the preheating plant, and the particular problems associated with the erection of such plants on working coke oven batteries. Commissioning and operating procedures are described and the problems associated with the integration of the preheating plant into the normal pushing schedule of the plant are discussed. Preliminary operating results with regard to the effect of preheating on coal throughput on the commercial plants and on the properties of the coke produced are given."

Jan 1, 1972

By M. B. Greyver

The results of laboratory tests on iron precipitation from non-decopperized nickel anolyte are discussed. The parameters (pH, oxygen consumption and stirring speed) for iron precipitation from the anolyte, using the oxygen in the air, were determined. In the experiments conducted under optimized conditions, the residual concentration of iron in the anolyte after precipitation was 0.7-2.7 mg/L, and the copper and nickel losses to the solid phase were 0.05-0.07% and 0.05-0.16%, respectively. The filtration ability of the iron slurry obtained was 2 to 5 times higher as compared to that after chlorine oxidation and oxyhydrolysis. The leaching of the non-ferrous metals from the iron cake made using the separate nickel anolyte purification process was possible by dissolution in sulphuric acid. It was established that after cobalt precipitation from the copper- and iron-depleted anolyte, the precipitates obtained had lower ratios of Fe:Co and Ni:Co.

Jan 1, 2006

By William J. Davis, R. Blair Hrabi, Sally Pehrsson, Amelia V. Logan, Ross Sherlock

"Abstract - The Meadowbank gold deposits are hosted by a series of polydeformed and metamorphosed auriferous iron formations located in the Neoarchean Woodburn Lake group, western Churchill province, Nunavut. The supracrustal rocks in the Meadowbank area consist of a thick sequence of intermediate-composition volcanic rocks with intercalated iron formation, ultramafic volcanic rocks, and quartzite to quartz arenite. U-Pb geochronology gives an interpreted age of 2711 ± 3 Ma for the host strata at Meadowbank.The intermediate volcanic rocks are predominately volcaniclastic and have a geochemical composition that is consistent with an active continental margin setting. Iron formation deposition was coeval with the intermediate volcanism, as indicated by incorporation of volcanic detritus in the chemical precipitate. The geochemistry of the iron formation indicates that it precipitated from a fluid that at one time was at high temperature (>250°C), reducing, and likely acidic. Intercalated ultramafic rocks rarely show spinifex textures and have trace element ratios that are consistent with mantle plume-related undepleted mantle asthenosphere, likely recording episodic mantle upwelling associated with rifting in an active continental margin. The contact between the ultramafic volcanic rocks and the overlying quartz arenite is a disconformity marked by a quartz pebble conglomerate. However, similar geology on the structural hanging wall and footwall of the contact suggests that the conglomerate does not represent a significant hiatus in deposition, and is more likely a prograding terriginous siliciclastic unit.Three metamorphic events are recognized. The first, M1, is a cryptic greenschist facies event that is pre-D2. The second event, M2, is a mid-greenschist to amphibolite facies, syn-D2 event. M3 is the last event recognized, and is a post-tectonic greenschist facies event that is regional but inhomogeneous in extent, possibly reflecting thermal aureoles around post-tectonic 1.7 to 1.8 Ga Nueltin granites. The structural geology of the area is complex with four phases of deformation recognized, two of which had a significant effect on the geometry of the deposit. All of these regional events are interpreted as Paleoproterozoic in age. Relationships between deformation fabrics and mineralization, as well as the overall geometry of the mineralized envelopes, suggest that the deposit formed during syn- to late-D2. Superimposed on the mineralization are D4 structural elements that postdate gold mineralization.The main control on gold mineralization is replacement of magnetite by pyrrhotite and pyrite in high-strain corridors. The composition of amphiboles and chlorites associated with gold mineralization is remarkably consistent and shows no spatial or temporal variation, suggesting that it was buffered by the iron-rich nature of the host rocks. The bulk metasomatic effect on the intermediate volcanic rocks is characterized by the addition of K2O and the loss of CaO, Na2O, and MgO, with little variation in the total iron content. The elemental losses are likely the result of destruction of feldspars, stilpnomelane, and ferroactinolite, which are relatively common in the unmineralized rocks, and the formation of sericite, chlorite, and grunerite."

Jan 1, 2004

By L. J. Cabri

A gold sample assaying 11.4 g Ault was selected for this study. Approximately 99% of the gold in this sample occurs as discrete gold minerals; the balance occurring as invisible gold in pyrite (determined by SIMS analysis). The sample was coarsely ground and sieved, yielding -38, 17, 22 and 22 wt.% in the -400+160 um, -160+80 um, -80+40 µm and -40 µm size fractions. Assays of the size fractions gave a gold distribution of -32, 20, 21, and 27% respectively. Processing with the hydroseparator (model HS-02) gave two tailings products and one concentrate from each size fraction. Assays of the concentrate from each size fraction gave gold recoveries of -7, 7, 24, 19%, respectively. Polished sections were prepared from each hydroseparation product and studied by image analysis to automatically search for gold minerals. Calculations were done using the surface area of the grains of gold minerals found on the polished sections.

Jan 1, 2005

By C. A. Pickles, K. Mackowiak

Black Thor chromite ore from the Ring of Fire deposit in Northern Ontario was carbothermically reduced in a microwave system to produce a ferrochromium alloy. The as-received ore contained 43.0% Cr2O3, 19.9% Fe2O3, 14.5% MgO, 13.4% Al2O3, and 7.0% SiO2. Activated charcoal was utilized as the reducing agent. Firstly, TGA/DTA and permittivity studies were carried out in order to determine the relationship between the reduction process and the microwave susceptibility. Secondly, microwave reduction experiments were performed on the chromite ore. The variables investigated were; the amount of carbon, processing time, input power and type of atmosphere. A maximum chromium grade of 76.1% and a chromium metallization of 82.7% were obtained for a processing time of 20 minutes, a power input of 800 W, and a carbon addition of 15% in an argon atmosphere.

Jan 1, 2017

By J. Aydt, W. McCombe, H. McLaughlin, L. Nightingale-Mercer

The McClean Lake mill is located approximately 850 kilometers north of Saskatoon, Saskatchewan, Canada and is operated by Orano Canada Inc. (Orano). The mill was designed to process high grade uranium ores, and in 2014 completed a multiyear expansion to increase its production capacity to 24 Mlbs U3O8 per annum. This expansion included significant changes to the mill’s secondary leach circuit in support of safely processing high grade ore directly from the Cigar Lake mine. In particular, inter-stage pumps were introduced for slurry transfer between the secondary leach tanks, replacing the original pressure feed forward operation. The design and selection process for these pumps was strongly influenced by the uncertainty related to actual radiation dose exposure when processing undiluted high grade ore (up to 30%U), and consequently the desire to maintain minimal occupancy in the vault enclosure containing the secondary leach tanks. These constraints resulted in the decision to install positive displacement pumps above the leach tanks in a suction-lift configuration. Operating experience with this design has shown that the pump capacity is significantly de-rated due to process solution gas-entrainment and pipe scale build-up, particularly in the front end of the secondary leach circuit where reagent dosage is high. Experience processing high grade ore also provided accurate radiation dose data to demonstrate that the level of occupancy in the leaching vault enclosure could be re-evaluated to accommodate alternate pump configurations. This paper reviews the lessons learned and improvements made to safely increase the capacity and reliability of the leach inter-stage pump system.

Jan 1, 2020

By D. T. Aikens, R. Gaudreault

"This paper summarizes structural repairs carried out on a marine wharf consisting of six 13.7 m by 13.7 m reinforced concrete cribs linked together by five 14.6 m long beam spans. As a result of problems during original construction and twenty years exposure to severe weather conditions, the deck support system between the cribs required major restoration work. In particular, some of the beam seats under the crane rail supports had marginal bearing area to support the structural loads. Faced with closure order for the wharf which is required to bring in raw materials to the plant and a long outage to make repairs as stipulated by Ports Canada (owners of the structure) alternate means of making the required repair were examined.The method chosen allowed the work to be done much more quickly, at lower cost, without taking the wharf completely out of service. The badly deteriorated seats were rebuilt using epoxy grout, tapered sole plates and a positioning jig."

Jan 1, 1988

By Hermann A. Derbuch

Twin Mining made a conscious decision when it changed from gold mine explorer and developer to become a diamond explorer in Canada?s Arctic. It was a decision of survival in an environment that saw junior gold companies either disappear or turn into dot com companies. Twin Mining learned fast and staked within weeks of that decision the first 50 sq. km of claims on, what became, its TORNGAT property at 59½° latitude on the east coast of the Ungava Bay in northern Quebec. This land position, 724 diamonds later (including 322 macro diamonds of gem quality, white and transparent) grew to 507 sq. km today. The 74° latitude called when Twin Mining exercised the option to acquire 3 claims owned by Helix Resources Inc. which revealed exceptional diamond potential. A gem quality 5.4 mm (0.768 carat white and transparent) diamond and the discovery of 16 pipes on 15.2 ha convinced Twin Mining?s board to sanction the deal. Both properties have outstanding features. Kimberlite Dykes and Pipes are in outcrop. The property locations have deep-sea access near existing infrastructure. Their locations are in an environment that is stable and development friendly and has a positive track record with successfully operating mines near-by.

May 1, 2001

By D. Auerswald, M. Kotze, V. Yahorava, T. Udayar

Uranium can be recovered from lower grade uranium ores (<800 mg/kg U3O8), leach residues, and waste dumps cost-effectively via resin-in-pulp (RIP). Using this technology, no solid/liquid separation or washing of the leach residues are required. Furthermore, the uranium recovery can be higher than with conventional methods. RIP technologies have, however, been perceived negatively due to concerns around potential resin fouling and losses. This paper highlights the advancements made in RIP and discusses a number of potential applications where the technology could deliver economic and consequential environmental benefits by reducing the uranium levels in the waste.

Jan 1, 2014

By Michael P. Davies

Mine waste facilities, and in particular tailings impoundments, are one of the highest risk aspects of any mining venture. There are far too many recent examples of how real and/or perceived physical or environmental impact from one of these facilities has had enormous economic impact on both the specific project and the overall parent mining company. In many cases, such dramatic impacts have followed a period where a due diligence of the mine had been carried out (e.g. in an acquisition, merger or similar) or an internal or otherwise audit had taken place. Consequently, it is suggested that there is a need for an audit/due diligence framework to assist those processes in terms of assessing mine waste facilities. This paper presents a recommended framework for completing either an audit or due diligence assessment of mine waste facilities under a number of typical scenarios. The framework will be of relevance to reviews carried out prior to mine construction, during operations and for mines that have been in a care and maintenance position and are to be re-opened. Concepts of risk management are introduced as an essential part of the recommended framework. The paper should be of value to current and prospective mine owners involved in stewarding an area of significant risk to viability of their mining venture. The paper should also interest prospective lenders to these ventures.

May 1, 2007

By W. G. Spencer, E. Sulis, P. M. Tiley

"Dome Mine&apos;s No.8 shaft is located about 1 km northeast of Dome&apos;s No.3 shaft and mill complex in South Porcupine, Ontario.No.8 shaft was brought into production in 1984 to hoist ore and waste for the complete mine as well as to service the mine from the 1600 level and below. Access above 1600 level is provided by No. 3 shaft.The rectangular shaft has six compartments consisting of two skips, a cage, counterweight, manway, and a service compartment. The over-all size of the shaft is 14 ft 7 in. by 20 ft by 7 in. (Fig. 1). A 7th compartment for shaft deepening is incorporated in the bottom 500 ft of shaft. The shaft is 5470 ft deep. The loading pocket is located at 5300 ft and the dump 61 ft below the collar.Surface facilities consist of a concrete headframe and a structural steel hoistroom which houses the cage and production hoists . The service hoist is a 12-ft diameter double drum, with one 1750 HP DC motor, and the production hoist is a 15-ft diameter double drum with twin 1750 HP DC motors.Both hoists have a single clutch and a single reduction gear train built by Orenstein & Koppel, and equipped with Canadian General Electric drives. The cage hoist is equipped with a disc pinion brake.The skips were originally aluminum 14-ton capacity, Rolla-chute design with steel liners on three sides and 4 in. of rubber on the back, to absorb the loading impact. The original guide rollers consisted of fixed 10 in. pneumatic tires. The face-to-face units were eventually replaced with larger airplane-type tires mounted on a spring loaded pivot arm."

Jan 1, 1993