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By Y. Faitakis, G. J. Powley, C. Mackenzie

"In the mining industry, maintenance is a key priority and a major expense. As mining equipment is large in scale, high in cost, and difficult to tow, it is key to maintain equipment on a timely basis. Breakdown in -40°C or away from a maintenance station can cause great costs in repair as well as lost production.It has often been quoted that about 5% of the North American production is lost every year due to unscheduled downtime. About onethird of the downtime is attributable to equipment failures. From this fact alone, it is evident that proper equipment maintenance can add tremendous value to the bottom line. As organizations and equipment become more sophisticated, maintenance expectations begin to evolve. This evolution can be traced through three generations:The first generation (1940–1950): A “fix it when it broke” attitude prevailed during that time, mostly because the industry was not highly automated.The second generation (1950–1980): Automation and mechanization of the industry increased. The industry now was dependent on machines performing the majority of the work. Preventive maintenance (overhauls at fixed intervals) appeared as the first maintenance practice. Maintenance costs started to rise rapidly.The third generation (1980–): Assets become more complex. Failure modes are better understood and computerized maintenance methods are becoming mainstream. Condition-based maintenance (CBM) starts to gain ground as a cost-effective alternative, replacing the fixed interval overhauls. As corporations amalgamate and market places are becoming more complex multi-functional teams become responsible for maintenance.At the beginning of the third generation, maintenance organizations began embracing technology to provide efficiencies in maintenance management. A strong push was made to transition from reactive maintenance to proactive maintenance. At the forefront of this push is condition-based maintenance."

Jan 1, 2004

By Ivan C. Machado

Jaguar Mining Inc. is implementing a 1.1 Mtpy, 7-year LOM gold project in the Iron Quadrangle (?Quadrilátero Ferrífero?), Brazil. Construction is on schedule and startup is programmed for Q2 2010. Mill feed grade averages 4.43 g/t and LOM gold production is estimated at 800,000 oz. The project entails two sulfide underground mines and three oxide/transition open pit mines. A state-of-the-art flotation/gravity concentration/CIP Plant has been designed based on testwork carried out by internationally recognized Canadian, U.S., and Brazilian laboratories. The mass recovery of the flotation section is only ca 10%, while its gold recovery is about 90%. Therefore, the CIP-ADR Plant will treat only 10% of the solids mass contained in the mill feed, while 90% will be fully cyanide-free material that will be partially used to feed a Backfill Plant. The leached tailings will be conveyed to a rather small tailings dam, whose 20- m3/h supernatant solution will be treated by a new electrolytic technology. The plant design contemplates a High Density Thickener for the cyanide-free rejects exceeding the backfill demand and a Regrinding Section before the Leaching Circuit. The estimated overall metallurgical recovery sits around 92%. This paper focus on the use of modern technologies to reduce costs, risks, and liabilities.

May 1, 2010

"10.1. IntroductionAs discussed by Westland (1981), the six platinum-group elements (PGE) belong to the Group VIII transition metals, to which iron, nickel and cobalt also belong. The chemical similarity between these nine metals accounts for the fact that they tend to be concentrated together as a result of geological processes. These processes started to effect the distribution of these metals at the very beginning of the earth's history.The first line in Table 1 gives the concentration of nickel and certain other Group VIII transition metals in carbonaceous chondrites, those meteorites with a composition that is believed to reflect best the composition of our planetary system apart from hydrogen and helium. The second line is the estimated concentration of the same elements in the earth as a whole. It is seen that these two estimates are quite similar, the earth being perhaps 2 to 3 times richer in these particular metals than the planetary system as a whole. The reason for this is not fully understood, but it may be due to the earth being formed by the accumulation and capture of a greater proportion of iron (and therefore siderophile-rich) than stoney (and therefore siderophile-poor) planetesimals than the planetary system itself.The third line in Table 1 is the average composition of the crust of the earth. With the exception of iron, all of the Group VIII transition elements are rare in the crust, present at concentrations of It tOO to Itt ,000 of that of the bulk earth. This is because the outstanding chemical characteristic of this group of elements is their siderophile nature; that is, their tendency to alloy with metallic iron. Because of this, once the earth became molten very early in its life history, these elements dissolved in the molten iron present at this stage and were removed from the outer accessible parts as this dense liquid sank to the centre of the earth to form the core. Although iron was the main instrument of concentration of the Group VIII metals, its much greater over-all abundance and its ability to combine with oxygen more easily than the others in the group means that significant amounts still remain in the oxidized portions of the earth.The fact that most of the Group VIII metals are available to use in workable concentrations at all is that they are 5 to 30 times more concentrated in the earth's mantle than the crust, as shown in line 4 of Table I. There has been a constant, if slow, interchange of matter between the mantle and the overlying crust throughout most of the earth's history and, under certain circumstances, this interchange has given rise to economic concentrations of metals."

Jan 1, 1979

By Wei Sun, Jaewon Choung, Zhenghe Xu, Zhangming Cui

"High intensity conditioning has been demonstrated in a number of cases to enhance fine particle flotation. The improvement could be attributed to the enhanced formation of particle aggregates. However, the understanding of aggregation mechanisms remains rather limited. Recently, an enhanced particle aggregation was observed by the formation of small air bubbles resulting from high intensity agitations (HIA). In this study, a novel visualization system is designed and constructed for characterizing the fine bubbles formed in an HIA cell. The new design allows determination of gas evolution and bubble size distribution. With this system, the effect of dissolved air on fine bubble generation by HIA is investigated. The results show that HIA creates a favorable environment for the formation of gas nuclei and small size bubbles due to hydrodynamic cavitation.INTRODUCTIONFlotation is a dominant separation technique employed by mineral industry to recover the valuable mineral particles. However, in general, particles finer than approximately 10 µm are not effectively recovered by conventional froth flotation technique due to their low inertia and high specific surface area (Fuerstenau, 1988). Therefore, the fine particle flotation has been considered as the central challenge of flotation research. Substantial advances have been made for fine particle flotation from both a theoretical and technological point of view (Fuerstenau et al., 1979; Song and Lu, 1990; Subrahmanyam and Forssberg, 1990; Sivamohan, 1990; Sedat and Barry, 1991; Matis, 1995). High intensity agitation (HIA), for example, has found some success to improve fine particle flotation kinetics in laboratory tests (Chen et al., 1999a and 1999b; Valderrama and Rubio, 1998; Rubio and De Brum, 1994).The effect of mechanical agitation on flotation kinetics has been recognized for many years. In mid-1970s, Warren (1975) reported that increasing the shear rate of the pulp could improve the efficiency of ultra fine particle flotation. He attributed this improvement in flotation to the formation of aggregates. This concept of aggregation has been extended by Rubio (1978), and Bulatovic and Salter (1989), who employed high intensity agitation (HIA) method to promote aggregation. They found that HIA with the addition of collectors, frothers and suitable modifiers significantly improved both the flotation recovery and selectivity of ultra fine sulphide minerals. The improvement was explained by warren’s shear flocculation theory. Till now, many studies (Chen et al., 1999a and 1999b; Valderrama and Rubio, 1998) have been undertaken to improve fine particle flotation by HIA. However, most of these studies were focused on the relation between energy input and flotation efficiency. Less information was available for describing how the aggregate formed during HIA process."

Jan 1, 2006

By W. P. Campbell

Foreword At the outset, it should be stated that this paper does not pretend to present a panacea for the ills that 'frozen' wells are heir to, but rather is an attempt to bring to the attention of those whose special concern it is, a little theory and some facts, which have been brought to light recently, in connection with the problem of 'freezing' as it applies to naphtha-producing wells in the Turner Valley field, and to draw some conclusions that follow from the facts discovered. It should also be stated that the investigation is not to be regarded as in any sense complete, but in progress. Doubtless new facts will be brought to light from time to time which may modify present ideas on the subject, but whether favourable or contrary, these will be welcomed, as the purpose of the Department is to secure as complete information as possible of the factors contributing to 'freezing', since this must be the first step to an intelligent attempt to remedy or prevent its occurrence, if such can be done. It is possible, too, that the cause or causes of freezing may not be the same in all cases, but whether this be so or not will have to be determined later as new information comes to hand. Introduction One of the worries which must constantly confront the operators of gas-naphtha producing wells in Turner Valley-although all, fortunately, have not met with it-is the possibility of the occurrence of what is commonly called 'freezing' (going off production).

Jan 1, 1930

By David M. R. Stone

Underground mining in the gold industry is gaining popularity in Nevada, and with it the use of cemented rockfill is becoming mainstream. Even with today?s high metal prices, however, mining companies are very sensitive to the margins they realize, and are constantly looking for ways to reduce costs or improve efficiencies. Given that the backfill is one of the largest cost centers in underground mining, mines are constantly under pressure to reduce backfill costs. This paper will present an overview of the factors that impact cemented rockfill quality in Nevada mines based on the authors many years of consulting experience. Cemented rockfill quality can be linked directly to binder consumption, hence programs aimed at improving quality generally result in lower binder costs, better performance in the fill (hence less dilution), and improved safety. The paper will conclude with the key elements of a comprehensive program of backfill quality assessments and backfill QA programs. Comparative laboratory and minesite data from a number of operating and proposed mines will be presented.

May 1, 2007

By Andre L. De Vegt, Cees J. N. Buisman

"For the past eight years Paques has been engaged in the development and installation of treatment systems based on biotechnological processes to remove sulfur compounds from water, air and gaseous streams. Metal and sulfate can be removed from mine waters using two biological steps:1. Sulfate reducing bacteria convert sulfate to hydrogen sulfide (H2S). The H2S reacts with the dissolved metals to form insoluble metal sulfide precipitates.2. Sulfide oxidizing bacteria convert excess H2S to elemental sulfur.Paques has designed and installed a groundwater treatment system for the Budelco zinc refinery in the Netherlands to remove metals and sulfate as described above. The system has been operating since May 1992 treating a flow of 5000 m3/d. A test program, utilizing a large scale pilot plant, will commence in November 1995 at a copper mine in the U.S.A. to remediate groundwater contaminated with sulfate and heavy metals. Biological reactor performance and selective metal recovery will be studied. Sulfur dioxide (S02) can be removed from fluegas using a combination of scrubber technology and the two biological steps. The end product is elemental sulfur. A commercial size biological desulfurization plant has been in operation since July 1994 treating 50,000 Nm3/h of fluegas. Sulfide can be removed from mining waters or anaerobic wetlands using a THIOPAQ® Sulfide Oxidizing reactor.lntroductionSulfur can be present in various water and gaseous streams in a mining operation. In many cases the sulfur compounds have to be removed because of toxicity, corrosive properties and contribution to acid rain. Mine waters can contain dissolved heavy metals. An efficient way to remove dissolved metals removal is the precipitation as metal sulfides. Paques has developed and installed treatment systems based on biotechnological processes to remove sulfur compounds from water, air and gaseous streams. These treatment systems biologically convert sulfate to hydrogen sulfide (H2S) which can be used in heavy metal precipitation. Excess H2S is biologically converted to elemental sulfur. This paper provides an overview of these newly developed biotechnological treatment systems."

Jan 1, 1997

By A. J. Hyett

Due to the ongoing revolution in electronics and communications more cost-effective geotechnical instruments can be manufactured using digital technology. This new generation of digital instruments have embedded microcontrollers that provide highly accurate and reliable measurement of geotechnical behaviour in real world units that can be immediately interpreted. Digital measurements can be transmitted across existing mine-wide wireless communication networks, interfaced directly with infra-red enabled PDA?s or PC?s, and parsed to memory by low-cost data-loggers.

Oct 1, 2004

By J. W. Spyker, D. H. Filson, W. H. W. Husband

"Large deposits of sodium sulphate exist in southern Saskatchewan in the brines or muddy crystal beds of alkali Jakes. Currently, more than 300,000 tons per year of sodium sulphate are produced from these lakes and sold as salt cake to the paper industry. Most of the production is obtained from the brines of a limited number of Jakes.The Saskatchewan Research Council was requested to develop a process for recovering sodium sulphate from the muddy crystal beds so that this type of deposit could be exploited in the future. A flowsheet was developed with the aid of phase diagram data and the process was studied on a pilot-plant scale using raw material from three different lakes. Satisfactory yields and product purities were obtained in all cases, and an economic analysis indicated that a 100,000-ton-per-year sodium sulphate plant could produce salt cake at a cost of $12.26 per ton for a rate of return upon investment of 16 per cent at current market prices."

Jan 1, 1966

By Matthew Violot

Introduction ?Located off highway 6, approximately 5 km, SW of Thompson, Manitoba ?Base metal mine: Nickel? Maximum Depth: 4100 ft ?Started Production in 1966 (INCO, 2006)?Halted in 1978?Re-opened in 1989 ?Two types of mining:?VBM (vertical block mining)?Cut and Fill 1.1 Problem ?Birchtree mine is experiencing failures and undesirable dilution in some of their stopes. ?Time can have a large effect on open stopestability. ?Birchtreeis interested in how time is affecting the stability of their stopes. ?The study focuses on the lower 84 complex, which extends from 2750 to 4100 foot level.

Nov 1, 2011

By Jean M. Richardson, Steven R. Williams

"Geometallurgical mapping is a new team-based approach that documents variability within an orebody and quantifies the impact of geology (host rocks, alteration and structure) and mineralogy on grinding, metallurgical response and metal recovery processes. The quantitative, spatially constrained database that results can be readily integrated into 3-D block models and mine planning activities. Thus it is an important tool to reduce the technical risk associated with new mine developments or expansions.Normally undertaken during pre-feasibility or feasibility planning stages, the first step of geometallurgical mapping is to review the geology, mineralogy and other critical parameters and construct a geometallurgical matrix. This matrix provides an objective base to then guide sampling and/or compositing activity for physical property and metallurgical testing. Depending on the project, an array of testing techniques can be used to characterize the ore and feed data into the geometallurgical model.While the range and number of tests needed is project-dependant, the geometallurgical mapping approach allows the development of empirical models or interdependent relationships, grounded by spatially-constrained real data. Extensive data sets collected using less expensive testing methods can be meaningfully correlated with results from complex specialty tests to yield realistic metallurgical recovery data.With geometallurgical mapping completed, the resulting database can be integrated into an overall 3-D block mine model. This means that metallurgical response can be used to realistically forecast the recoveries of specific blocks, perhaps sampled only by drilling. Mine planning can then incorporate the forecasted metallurgical response and generate future project cash-flows; opening the door to economic optimization of the mine exploitation."

Jan 1, 2004

By T. Belem

The main objective of this study is to better understand the effect of underground CPB placement conditions on its microstructural, hydro-mechanical and geochemical properties. To reach this objective, three PVC columns, each 3 m high, were built, and were filled with CPB material. The columns allow simulating undrained (UD), half-drained (HD) and fully-drained (FD) conditions along with the measurement of resulting self-weight consolidation settlement of the CPB. Bleed and drainage waters were also collected from each column for geochemical analyses (pH, Eh, EC, SO42- content). After 90-days of curing time, a total of sixty CPB samples were cored from the backfill columns and subjected to uniaxial compression tests followed by physical characterization (Gs, void ratio, water content, degree of saturation, specific surface area). The results showed that the total drainage water and the maximum observed settlement occur mainly within the first 72 hours after the columns are filled. The corresponding volumetric strains suggest that in situ backfilled stopes simulated behave in a similar way to the fully-drained (FD) or the half-drained (HD) conditions. The MIP test results show that the pore structure varies between the top and the bottom of each column.

May 1, 2007

By Paul J. Sylvester, Yeonuk Choi, Mike Nelson, Louis J. Cabri, Mike Tubrett

"The development and characterization of sulphide standards containing as many as seven precious metals has increased the efficiency and comprehensiveness of micro-analyses for deportment studies using laser ablation microprobe inductively coupled plasma mass spectrometry (LAM-ICPMS). Because trace precious metals occurring in solid solution or as sub-micron inclusions in sulphides, arsenides, and sulpharsenides are often heterogeneously distributed, a significantly large number of analyses are required. Using well-prepared samples it is possible for up to 120 analyses per day, far more than by other comparable methods. Examples will be discussed from projects focused on the deportment of Au and the platinum group elements (PGE).INTRODUCTIONBecause gold and the platinum-group elements (PGE) occur as both discrete minerals and in trace quantities in a variety of minerals (mostly sulphide, arsenide, and sulpharsenide minerals), specialized trace analytical methods were developed and applied to determine the deportment of precious metals in ore deposits, requiring sensitivities ranging originally from about hundreds to presently tens of ppb for individual precious metal or less.In mineralogical studies, micro-analytical techniques such as the electron microprobe were first applied to obtain analyses for major elements in about 1959-60, but as the instrumentation improved early attempts for PGE in potential carrier minerals indicated that reasonable levels of detection (LOD) were of the order of 300-400 ppm, depending on analytical conditions, X-ray spectral lines, and mineral matrix (e.g., Cabri and Laflamme, 1976). Eventually, the lowest detection levels by electron probe micro-analyses (EPMA) obtained for some PGE were of the order of 10s of ppm using high accelerating voltage and current for extended measuring time on peak and background (Gervilla, Cabri, Kojonen, Oberthür, Weiser, Johanson, Sie, Campbell, Teesdale and Laflamme, 2004)."

Jan 1, 2010

By D. M. Wyslouzil

"Micron or sub-micron size gold inclusions in sulphide minerals cannot be extracted by conventional cyanidation. The gold surfaces must first be exposed by decomposition of the sulphides. As an alternative to roasting, pressure oxidation was found to be a most satisfactory method to decompose the sulphides and expose the gold particles. For cyanidation the pressure oxidation pulp has to be neutralized and stabilized with lime. It was found that thiourea can extract the gold directly from the pressure oxidation pulp without pH adjustment and oxidant addition. The main obstacles against the common use of thiourea are circumvented by a recently described control method of the emf during leaching.INTRODUCTION Amalgamation and cyanidation are the most conventional processes for gold extraction. The prerequisite for gold extraction is the possibility of gold surface to be contacted by mercury or by the cyanide solution. Encapsulated gold has little chance of contact and this leads to poor extractions.The reasons why a gold ore may be refractory are: micron or sub-micron size gold particles included in the sulphide minerals, such as pyrite and arsenopyrite; presence of and association with antimony, arsenic, tellurium and carbonaceous minerals. Some alternative processes, such as double oxidation(1,2), pretreatment with lime , H202, chlorine and Caro's acid, roasting and thiourea leach, have solved some of the problems of ""refractory"", but the sub-micron gold in sulphide minerals still shows little response. Roasting and the cyanidation of the calcine was one of the most successful approaches for this type of gold ore.There are two disadvantages to the roasting process: (1) the generation of sulphur, arsenic, antimony and mercury containing gases is undesirable and the control is extremely expensive; (2) gold particles might be fused in the roasted product and the oxidized base metals in the calcine might reduce the cyanidation efficiency. In recent years the pressure oxidation process has been considered as one of the more effective and satisfactory methods to decompose the sulphide minerals and expose the gold surface to the cyanide leach solution. The pollution problem could be better controlled and the gold extraction is more amenable."

Jan 1, 1985

By Robert B. Parsons

"Canada's mining industry has reason to celebrate the federal government's budget which was tabled on February 26, 1986. The budget contains an important proposal to improve the flowthrough share rules by effectively eliminating the investor's liability to third parties. This proposed rule change and other important budget proposals are the subject of this month's edition of Tax Notes. IntroductionWhile Finance Minister Michael Wilson's February 26, 1986 budget received a mixed reaction from business and labour leaders, the budget was mainly ""good news"" for the mining industry. A long-standing bugaboo in the flow-through share rules will be cleared up, a permanent set of rules will be installed for the tax treatment of Northern allowances, and limited partnerships will apparently be able to continue to serve a useful role in financing exploration programs. Flow-Through SharesThe present rules which apply to an investment in flow-through shares are structured in such a way that the relevant Canadian exploration expense (""CEE"") must be incurred by the investor, and not by the issuing corporation. Obviously, in the case of most flow-through share issues, this requirement that the investor incur the CEE poses a practical problem-the investor typically does not have the expertise or organization to incur the CEE. As a result, an integral part of most flowthrough arrangements is an agreement whereby the investor appoints another party (usually the issuing corporation) as his agent to spend the flow-through funds on CEE on his behalf. Although the issuing corporation might consequently be the party which actually spends the flow-through share funds on CEE, nevertheless the legal structure is such that the investor is the ""principal"" who is incurring the CEE. As such, the investor is legally responsible for any third-party claims with respect to the exploration program. Third-party liability might arise, for example, in the case of property damage, personal injury, environmental infractions, and so on."

Jan 1, 1986

By R. W. Lawrence, J. B. Austin, P. B. Marchant, J. T. Chapman

"Biohydrometallurgy is the use of microbiological reactions to assist in the extraction of metals. Bioleaching offers an innovative alternative for the treatment of refractory gold-bearing sulphides (1). One particular cost advantage of bioleaching is the potential to enhance gold extraction after only partial sulphide oxidation. The apparent selective sulphide oxidation is more prevalent where the refractory gold is associated with arsenic and antimony (2). Oxidation of the pyrite minerals and arsenopyrite at different oxidation rates can result in high gold extractions at relatively low levels of sulphide oxidation due to preferential, or selective, oxidation of the arsenic mineral. Arsenic extraction rates in a bioleach process typically exceed the iron extraction rate. Therefore, in concentrates in which the precious metals are principally associated with arsenic minerals, liberation of the precious metals may be possible with high recoveries and less than complete sulphide oxidation. The capital and operating cost advantages of partial sulphide oxidation are inherently obvious and are briefly illustrated herein.Ultra fine milling (UFM) is another alternative for enhancing gold extraction from refractory sulphides. Some refractory gold is simply particle size sensitive and UFM can improve gold extraction without the need for costly sulphide oxidation (3). Fine grinding technology for enhanced gold recovery developed by Metprotech, South Africa, is currently producing 2 - 4 micron ground product using a single pass, stirred ball mill. Proper application engineering has ensured that plug flow occurs through the mill. Plug flow from the feed inlet at the bottom of the reactor ensures that short circuiting of feed does not occur and as a result, the mill product does not require classification.A combination of UFM as a pretreatment for bioleaching could provide a mechanism for enhanced selective sulphide biooxidation and/or increased overall gold extraction following biooxidation due to improved particle liberation ( 4). The effect of UFM on biooxidation costs will be discussed. Only sulphide concentrates are considered herein for application of UFM in combination with biooxidation."

Jan 1, 1990

By Renaud Adams, Vladimir Salatic

"The El Mochito Mine located in Honduras, Central America, has produced lead, zinc and silver in concentrates since the 1950s. The paper discusses changes made in both lead and zinc circuits to reduce costs while maintaining production during the recent period of low metal prices.The intent of this paper is to show that sometimes thinking “outside the box” is beneficial for progress. Also, in striving for better metallurgical results we should not forget the basics of flotation. Sometimes the latest trends may not be applicable to the ore we are treating and may not meet the operating cost constraints.A major beneficial impact on El Mochito metallurgy was achieved when the overall grind was coarsened, eliminating the significant percentage of slimes. Switching from two parallel twostage grinding circuits to one circuit with three-stage series grinding has resulted in a significant reduction in power and steel consumption. With minor layout/reagent modifications, both lead and zinc circuits gained not only in retention time but as well in producing better grade products.Of course, the challenges are still there; they exist and only with the full support of operational data collection and metallurgical laboratory investigation success can come. That is why, we as professionals, can never stop in trying to make the most of what we have.INTRODUCTIONThe El Mochito Mine is owned by American Pacific Honduras S.A. de C.V. (AMPAC) a subsidiary of Breakwater Resources Ltd. Of Toronto, Canada. The mine is located in western Honduras, approximately 220 kilometers northwest of the capital city, Tegucigalpa and 90 kilometers south of the country’s second largest city, San Pedro Sula. The mine began operating in 1948 and to date has produced more than 19 million tonnes grading 240 g/t Ag, 3.5% Pb and 7.4% Zn. Production in 2005 was 700,190 tonnes grading 87 g/t Ag, 1.85% Pb and 6.61% Zn.Decreasing metal content in the ore followed by the lowest zinc price on the market in the last 15 years (see Figure 1) has reduced the cash flow and endangered to stop the operation."

Jan 1, 2007

By Deon van Tonder, Chuck Edwards

"Solvent extraction is used in many acid leach uranium circuits to upgrade and purify the uranium bearing solution prior to final product precipitation. A number of strip options are available to recover the uranium from the loaded organic. This paper compares some of these options on the basis of typical capital and operating costs for both the solvent extraction circuit and downstream uranium processing, and also considers other related issues. The aim is to provide an objective comparative assessment of these options that could be used as a qualitative guide in process selection or improvement.INTRODUCTIONIn the 1950s the uranium industry was the first industry to adopt solvent extraction (SX). Since then, the flow sheets have advanced. Notable was the introduction of tertiary amine extractants by General Mills during the late 1950s and 1960s (Mackenzie, 1997; Mackenzie, 2005), which is still today considered ‘state of the art’ technology. Initially solvent extraction was done in combination with ion-exchange, such as in the Eluex flow sheet, followed later by the adoption of direct solvent extraction from pregnant leach solution (PLS). From a strip-precipitation perspective ammonia strip followed by ammonium diuranate precipitation was widely adopted as the norm in the uranium industry. More recently alternative strip systems have been receiving increasing attention due to the environmental, health and safety issues associated with ammonia and the resulting calcined uranium oxide product. This paper explores some of these alternative options in comparison with the ammonia route.It is important to acknowledge that each operation is unique, and therefore the information and philosophy outlined in this paper serve to provide a general guide that can be potentially applied and adapted to any site or project."

Jan 1, 2013

By Ken Dyck Sr.

"Construction of the Cigar Lake uranium mine in northern Saskatchewan is well under way. Production is anticipated to begin in the first half of 2007. When in full production, the mine is projected to produce a net 18,000,000 pounds of U3O8 per annum from a high grade ore averaging approximately 20% U3O8.Water management was a major issue that required substantial effort in the Cigar Lake design. Not only is effluent quality controlled, a tight water balance to minimize effluent volume is also designed into processes in order to minimize environmental impact. This paper focuses on the water treatment process and water handling equipment, as well as the operating philosophies that will reduce fresh water consumption to a minimum. An overview of the mining and ore handling systems, as well as aspects of controlling water production from the ore formation and the mine development areas will also be discussed.INTRODUCTIONThe Cigar Lake ore body has proven and probable reserves that total 232 million pounds U3O8. Cigar Lake is the world’s largest undeveloped high-grade deposit. It is the second largest highgrade deposit surpassed only by the McArthur River deposit another Cameco - operated mine in the Saskatchewan Athabasca Basin. Cigar Lake is owned 50% by Cameco Corporation, 37% by AREVA (COGEMA Resources Inc.), 8% by Idemitsu Uranium Exploration Canada Ltd. and 5% by TEPCO Resources Inc. When production starts in 2007, ground ore slurry from Cigar Lake will be shipped to the McClean Lake mill (owned by AREVA) some 80 km away. All of the Cigar Lake uranium will be leached from the ore at McClean Lake. During initial production ramp up, all uranium will be processed at McClean Lake to yellow cake (uranium oxide). After the average production has exceeded a 10 million pound per year rate, Cigar Lake uranium will begin to flow from the McClean Lake Mill to the Rabbit Lake Mill. At this point, subject to regulatory approval, 57% of the Cigar Lake uranium, in a uranium rich solution, will be transported to Rabbit Lake mill (owned by Cameco) for purification to yellow cake. A complete summary of the flow of uranium from the Cigar Lake ore body through to product at both McClean Lake and Rabbit Lake can be found in a recent paper titled “The Cigar Lake Project – Mining, Ore Handling and Milling” in the March 2004 issue of the CIM Bulletin. The paper was written by C.R. Edwards."

Jan 1, 2006

By James A. Brown, Marius Botha, Christopher A. Fleming

"The technical and economic feasibility or re-processing hundreds of millions of tons of tailings in a number of historic gold tailings dams in the Witwatersrand area of South Africa is being investigated. The objective is to recover uranium and residual gold from the tailings, as well as to oxidize the sulphides to eliminate long term environmental liabilities associated with acid mine drainage. The tailings from the re-processing plant will be re-deposited in a new area far from urbanization, allowing for the rehabilitation of the real estate currently occupied by these historic tailings.The first step in the process involves a mild-regrind and mild acid leach of the tailings (pH 2.5 to 3.0) to dissolve up to 40% of the uranium. This uranium is recovered directly from the pulp by resin in pulp (RIP) with a strong base resin. The RIP tails are then treated by flotation to recover a rougher pyrite concentrate, and flotation tailings that are discarded. The rougher concentrate is cleaned to produce a final pyrite concentrate and cleaner scavenger tailings, and the pyrite concentrate is oxidized in an autoclave to leach uranium and generate acid. The autoclave discharge is combined with the cleaner scavenger tails, and the acid and ferric ions generated in the autoclave are used to leach more uranium and oxidize more sulphide in the cleaner tails. The product from the concentrate leach circuit is processed by resin in pulp for uranium recovery, followed by neutralization, cyanidation and carbon in pulp for gold recovery.The rougher float tails may be further treated to recover gold that was not recovered in flotation, but which had been liberated from the old dumps by natural oxidation of some of the pyrite over many decades. The float tails would be neutralized and treated by cyanidation and CIL to recover this liberated gold.All parts of the flowsheet have been extensively tested in the laboratory, and the critical elements of flotation, pressure oxidation and resin in pulp have been demonstrated in pilot plants. The results of the hydrometallurgical investigations are presented here."

Jan 1, 2010