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By H. Kuno, H. Kobayakawa

A borehole inclinometer with FBG optical fiber sensors has been developed to clarify very slow slope movements, and the results of its application to a real slope have been shown. The main characteristics kept in mind while developing this inclinometer are as follows. Firstly, the workability must be good so that only one cable is required for taking the measurements, because the inclinometer is composed of several FBG sensor elements constructed on an optical fiber. Secondly, the minimum interval of the distorted measurements must be close, namely, within 100 mm. Thirdly, the inclinometer must excel in thunder resistance. The measurements were executed by setting up the developed borehole inclinometer on a slope where the rate of displacement is several millimeters per year. The movement of the device was developed by comparing it with that of the results of an existing borehole inclinometer. The area of the displacement concentration, 1 m in width for 4-m sections, was confirmed by the existing borehole inclinometer. The boring core, recovered using a high-quality drilling technique, was examined. Landslide clay was not permitted at the location where the displacement had concentrated as a result of comparing the measurement results with the boring core, and the rotation of gravel by the movement of the slope’s land block was allowed. These results show the slope being caught in the stage just before a primary landslide. In addition, the feature whereby the slope behavior depends on the depth was analyzed by comparing the information on the underground displacement with precipitation and earthquake records.

Jan 1, 2015

By Richard Phiri, Mike Milupi, Jackson Kapobe, Anton Lemmer, Charles Mazala

Sustainability of plant operations in a rapidly changing mining environment requires a well-trained workforce to ensure smooth running of operations. To ensure availability of trained labor, Mopani Copper Mines PLC set up a training facility in 2013. The school offers training in Engineering, Mining and Process. The engineering section enrolls grade 12 school leavers for artisan training. Mining section offers comprehensive safety induction program to employees and contractors to meet the Mopani zero tolerance accidents safety policy. The process training section offers advanced training programs for plant operating personnel and university graduates. The program includes needs analysis, theoretical metallurgical modules, practical plant training and e-learning modules. The school also offers assaying and sampling courses. All the above courses are extended to individuals and other mining houses outside Mopani wishing to have their personnel trained. The Mopani sustainable development academy ensures that contractors engaged on the Mopani sites are fully registered, comply with all relevant regulations and have undergone induction program before they can be allowed to work on the mine. INTRODUCTION “Sustainable operation of the mining industry requires a continuous supply of trained human resources” (Kapobe et al., 2018). The publication by the Asian Development Bank (2014) argued that there was a mismatch between the demand for high-level skills and the oversupply of low-level skills. Globally, the mining industry requires continued supply of skilled manpower for its complex operations and long term projects whose sustainability is absolutely necessary. Implementation of newer technologies in plant operations to ensure cost effective and efficient operations requires trained and skilled labor at various levels to handle ever-changing challenges. Dutz (2007:152–156) made three recommendations for Building Worker and Management Skills in India: Strengthen enterprise - based training. The government should help ensure that the benefits of in- service training are widely recognized by enterprises, while simultaneously providing strong financial incentives Improve vocational training - Vocational education and training systems have been unsuccessful in preparing graduates to meet market needs, particularly because of a lack of interaction with industry. Public and private vocational education and training providers should be strengthened to make them more responsive to industry demands. Curricula should be updated to reflect modern technologies. Strengthen business, management education, management of research and commercialization - business and management education should be strengthened through public-private partnerships. Innovative management courses should be introduced. Although these recommendations were made for India, they also apply to Zambia and other low and medium income nations.

Jan 1, 2019

By George Hope, Brigitte Lacouture

"The Red Dog Mine has increased mill throughput for most years since startup. Although flotation equipment has been added over the last 10 years, the Red Dog metallurgical performance has been restricted by a lack of adequate flotation capacity. Two reduced tonnage trials conducted in 1997 as well as numerous SAG relines have consistently indicated the potential to significantly increase both zinc concentrate recovery and grade with additional flotation capacity. Late in 1999, a new circuit was designed with increased flotation capacity while simplifylng the current circuit, providing for three stages of dilution cleaning, and creating a retreat section. The purpose of the retreat section is to target the more complex particles and allow the chemical environment to be fine tuned for improved iron depression. In total, twentyone (21) Outokumpu OK50 mechanical flotation cells and four (4) 3.66m dia. column cells will be added to the current flotation circuit. The VIP (Value Improvement Program) Mill Optimization project is also addressing thickening, water treatment and power issues and has a total budget of $US105 M. Startup of equipment will be phased in to minimize the impact on current operations and is expected to be complete by the end of Q3-2001.INTRODUCTIONRed Dog Operations started production in November 1989. The mine is located above the arctic circle in northwest Alaska. It is linked by an 80 km road to a port-site facility where the concentrate is stored. Concentrate is shipped to markets throughout the world during the one hundred (100) days of open seas. Typical of sedimentary exhalative deposits, the metallurgy proved to be challenging and coupled with many mechanical shortcomings, the initial few years were very difficult. The start-up issues were slowly resolved and in 1998, a major grinding expansion, the PRI (Production Rate Increase) project, was completed. The project was designed to allow a 40% increase over the 1995 throughput. In 2000 mill throughput was 3,050,000 tonnes to produce a total of 960,000 and 140,000 dry tonnes of zinc and lead concentrates respectively, making Red Dog the largest zinc concentrate producer in the world. To address a shortage of flotation capacity remaining from the PRI grinding expansion, a project designated as the VIP (Value Improvement Program) Mill Optimization project was initiated. The mill optimization project is targeted to produce 1,100,000 tonnes of zinc concentrate at 56% zinc grade and 87% zinc recovery. The project will add twenty-one (21) Outokumpu OK50 (50m3) flotation cells, four (4) 3.66mO column cells and address water treatment, thickening and power generation capacities."

Jan 1, 2002

By Norman O. Lotter, Dominic Fragomeni, Jorge Oliveira, Dee J. Bradshaw, Rebecca Monnapula

"The role of mixed collectors in enhancing the metallurgical performance of sulphide flotation processes was reviewed in 2008, and later published as an integrated theory by Lotter and Bradshaw, 2009. The theory proposed that xanthates, dithiophosphates and dithiocarbamates could be synergistically combined in an optimal formula to deliver improved grade and recovery. The approach uses a five-step structure, viz.1. A mineralogical study of the ore to be treated, identifying and quantifying the economic minerals (with hosting and association) and gangue species present.2. Selection of candidate collectors from this mineralogical information and from the known pulp conditions of that operation (e.g. Eh, pH).3. Design and execution of a factorial experiment to estimate and test the main effects and interactions, and to model the response surface for an optimum formulation [Box et al., 1978].4. Use of a High-Confidence Flotation Test platform in the factorial to reduce overall errors and allow for clearer observations [Lotter, 1995; Lotter and Fragomeni, 2010].5. Statistical design of a plant scale trial to measure and validate the gains [Napier-Munn, 1995; Lotter et al., 2009]. Eland Platinum, which was commissioned in November 2007, appointed a team to apply this theory to formulate an optimized mixed collector suite for use in the operation. The West Pit ore (UG2 - Normal ore) was sampled at site and brought to Xstrata Process Support, Sudbury, Canada, for mineralogical and metallurgical testing in 2008. Microprobe and QEMSCAN measurement characterized the type, grain size and textural association of the discrete Platinum Group Minerals (PGM) present. From the mineralogical data, it was clear that a mixture of dithiocarbamate and two xanthates would be a likely reagent suite selection. A factorially designed set of High-Confidence Flotation Tests studying these was performed, concluding that an optimal suite using certain amounts of these collectors was a viable formulation that would advance the grade and recovery performance at Eland above the standard PIBX by approximately 2.13% 4E PGE. The plant trial, which used a form of the replicated block design after Napier-Munn, 1995, and Lotter et al., 2009, exceeded this estimate with an overall PGE recovery gain of 2.48% (absolute basis) and a concentrate grade gain of 26.4 g/t PGE (absolute basis), as well as reducing chrome in concentrate from 2.04 to 1.62% (absolute basis). These gains proved to be statistically significant using the analysis of variance method, and have validated the theory."

Jan 1, 2011

By E. Amini

"While methods exist to allow the simulation and optimization of existing flotation operations, there is a deficit in the ability to predict the performance of a new ore type pre-processing. A laboratory scale batch flotation testing methodology has been developed to assess the flotation potential of ores (known as the JK Floatability Index or JKFI). Previous work conducted by Alexander and Collins (2008) showed that the test could be used to predict full scale results in base metal ores. In this work, JKFI tests were applied to three different types of ore and one blend from a greenfields gold deposit to show the test applicability to flotation of precious metals. The results from the JKFI test on one ore type were then used to predict the response of all ore types (including the blend). To validate the predicted kinetic parameters, a comprehensive survey was performed on a pilot plant where each of ore types was processed in the circuit. An acceptable agreement was found between the experimental recoveries and predicted recoveries which indicate the potential of this method for use in predicting the flotation response of an ore.INTRODUCTIONOver 100 years, flotation has been considered an efficient method to concentrate different types of sulphide and oxide minerals and coal. Thus, numerous research studies have been undertaken to find an appropriate model to predict the behaviour of particles in flotation circuits (Barnwal et. al., 2006; Bascur and Herbst, 1983; Cecile and Villeneuve, 1991; Deng et. al., 1996; Zheng et. al., 2006). Most methods are based on the ore flotation response in an industrial plant which is expensive and time consuming to characterise. It is difficult to reliably predict flotation response for new types of ores in industrial scale plants.The floatability component modelling (FCM) approach, developed within the Australian Minerals Industry Research Association (AMIRA) P9 project, has been used to derive models of operating flotation plants for the purpose of simulation and optimisation (Alexander and Morrison, 1998; Alexander et. al., 2000; Varadi, 2006). This approach has been applied at a large number of operating sites treating a wide variety of ores. The approach requires metallurgical surveys to be undertaken in conjunction with batch flotation tests and cell characterisation. This allows ore and machine parameters to be derived, which are then used in the P9 Flotation Model. The development of a FCM generally requires a large commitment of resources and time. While there has been significant support of this approach in industry, the FCM approach is limited to ores already being processed (Runge, 2004; Wightman, 2006)."

Jan 1, 2009

By A. C. Wilson

In 1991, local prospector C. ?Mickey? Clement recovered at least three alluvial diamonds from the Michipicoten River. Two of these stones were sent to the Department of Mineralogy, Royal Ontario Museum, where they were identified as industrial grade diamonds with weights of 1.05 and 1.13 carats. Both stones were described as frosted and graphite-inclusion riddled. In 1995, prospector Sandor Surmacz and geologist Marcelle Hauseux discovered diamonds in bedrock approximately 20 km north of the Michipicoten River. An 18.1-kg bulk sample of a xenolith-rich lamprophyre, taken from an outcrop at the side of the Trans Canada Highway north of Wawa, yielded 1 macro diamond and 5 micro diamonds. All but one was gem quality (Sage 2000). Since then, more than 36 occurrences of diamonds in bedrock have been reported in an approximately 30 km2 area centred 20 km north of the town of Wawa. The occurrences are hosted within a sequence of unusual, Archean-aged, heterolithic breccias previously considered to have no economic significance. Ten kimberlite occurrences also have been confirmed in the Michipicoten greenstone belt in the last 5 years. Two of these kimberlites are diamondiferous. To date, expenditures on property acquisition and diamond exploration in the Wawa are estimated at $120 million (CAN). This paper summarizes the geological observations made on these occurrences and discusses including the relationship between the occurrences and the regional geology.

May 1, 2003

By W. Bullen

A hypothetical economic model has been developed for diamond projects in the Canadian Arctic using data from recent exploration and mining projects. The model reflects economic conditions applicable to the diamond mining industry in the Canadian Arctic at the present time. The economic model captures pre-production expenditures, production data, and post-production costs. Before- and after-tax cash flows are derived from the data along with the relevant economic criteria. From the corporate perspective, the model provides guidance concerning the minimum size and/or in situ value per tonne requirements of targets being sought in order to ensure economic success and predict exploration and development expenditures. From the government perspective, the model provides an indication of income and mining tax revenues that can be expected from a particular diamond project. The model shows, on average, diamond projects in the Canadian Arctic are expensive to explore for and develop, but are economically robust under the present conditions. They generate great economic return for mining companies and significant amounts of tax revenue for governments. The fact that diamonds can be transported directly from the minesite by air makes them particularly attractive targets for mining companies that operate in the Canadian Arctic.

Jan 1, 2006

By Goutam H. Shahani

Today, the mining and mineral sectors are being impacted by highly fluctuating supply ? demand, extreme price volatility and tremendous market uncertainty on a world-wide basis. This pattern is expected to continue for the foreseeable future. In particular, large consumers of tonnage quantities of industrial gases (oxygen, nitrogen) all have to make informed investment decisions in this uncertain economic environment. Plant owners ? operators need to carefully examine their current and future industrial gas needs in order to develop the most cost-effective industrial gas plant configuration. This can be done by partnering with an experienced engineering company that offers a complete technology portfolio for producing industrial gases. Ideally, the mineral and engineering companies have to work together as a single team to identify the most efficient and economic plant design. This includes an in-depth assessment of capital and operating costs, in order to deliver an optimal solution, taking plant reliability and process safety into account. It is important to consider all the industrial gas needs at a given manufacturing complex, in conjunction with available utilities and infrastructure to develop the best possible long term solution. Oxygen (O2) requirements include: oxygen enriched and oxy-fuel combustion, smelting and hydrometallurgical operations. Nitrogen (N2) is used for annealing. All of these industrial gas requirements have to be examined at a given site over a long term horizon in a holistic manner. Cryogenic air separation was first practiced at a commercial scale in 1902. Since then, the technology has been constantly refined to reduce cost; and improve reliability and flexibility. Today, using advanced engineering techniques, including proprietary process simulation, and modern fabrication techniques, it is possible to build air separation plants that are both low cost and high quality at the same time. It is important to work with an engineering company that owns a complete portfolio of industrial gas technologies, functional engineering capabilities and a variety of plant solutions. Such a partnership can quickly identify a ?sweet spot? for a given situation. By drawing on experience gained from hundreds of projects, it is possible to quickly optimize a given requirement. In situations where a customer wants a complete solution it may be appropriate to provide an air separation plant on a turn-key basis. This will enable the mining and mineral processor to focus on their core business. By using global procurement and construction capabilities it is possible reduce cost and improve project execution, thereby making industrial gas plant procurement simpler and cheaper for the mining and mineral sector.

Nov 1, 2011

By Norman O. Lotter, Simon Yu, Tony Deng, Antonio Di Feo

"Xstrata Nickel’s Raglan operation, northern Québec, achieved paymetal recovery gains in 2007 by switching the collector from potassium amyl xanthate to sodium isobutyl xanthate. To follow on from that position, Xstrata Process Support (XPS) and A. Di Feo of Raglan proposed that a mixed xanthate system be formulated and tested towards further performance gains. This work selected ethyl and isobutyl xanthates as the candidates. An optimized ratio of these was to be determined by a designed experiment. This suggestion was based on an abundance of published literature that reported mixed collector systems outperformed single ones. A partial factorial design was used to configure four test conditions and a mid-square for the flotation work. The testwork was conducted using High Confidence Flotation Testing. The raw data were analyzed and modelled by Design-Expert software. An optimum ratio of 1.6:1.0 by mass (1.32:1.0 in molar terms) of isobutyl:ethyl xanthate was determined. Without this modelling, the optimum ratio would have been missed. Results obtained with the new baseline flotation test using this optimized mixture showed a 3.3% Ni recovery gain in the bypass concentrate and more than 1% recovery gain for PGM over the single PIBX system. This system has been proposed for and accepted as a plant scale trial in the first half of 2010.INTRODUCTIONConcentrator OperationsMilling and flotation of sulphide ores ahead of smelting and refining is a common practice. This choice results in the larger and more variable paymetal loss in the production of saleable metal [Cramer, 2001]. For this reason, it is common to see an ongoing performance improvement programme at the operations level. When successfully performed, these e"

Jan 1, 2010

By R. P. D. Graham

It is almost exactly two hundred years since the foundations of the mining and metallurgical industries in Canada were laid. There '."'as nothing spectacular about this early start. It had to do with the most basic phases of these industries-the mining of coal and iron ore, and the manufacture of cast and wrought iron. At several points along the coast of Cape Breton island, coal seams may be seen outcropping quite conspicuously, and it is probable that they had been observed, and possibly worked in a small way, in very early times. However that may be, these coal seams have the distinction of being the first in North America of which there is any printed record, reference to them appearing in a volume by Nicolas Denys which was published in Paris in 1672. Regular coal mining in the district did not commence until about fifty years later, however. The metallurgists started their operations along the north shore of the St. Lawrence river. Here, at many points between Montreal and Quebec, are to be found deposits of bog iron ore, and in 1 730 a furnace for smelting such ore was established on the St. Maurice river. The St. Maurice forges continued in operation until 1880, and throughout that long period they were famed for the fine quality of castings produced, and also for a-superior wrought iron.

Jan 1, 1924

By P. P. Sinha, D. N. Bhatia, B. R. Ghosh, R. K. Gupta

Medium carbon, low alloy, high-strength steels containing Ni, Cr, and Mo are used for the fabrication of various types of high-strength fasteners required for launch vehicle structures. Cold heading properties and response to hardening and tempering are the essential requirements of the material for fabrication and application, respectively. Combining these two requirements is essentially warranted in high-strength fastener-grade steels. Cold heading properties require ductility, which can be achieved through spheroidizing treatments. In the absence of a suitable spheroidizing cycle for medium carbon 3.8 Ni, 1.8 Cr, 0.3 Mo steel, experiments were carried out with a theoretical understanding of the alloy system. This paper presents the experiments and results obtained after a spheroidizing treatment, which was carried out to optimize the process. Hardness measurements were carried out and microstructural studies were conducted to assess the spheroidization process and to determine the optimal spheroidization parameters.

Jan 1, 2007

By Glenn Norris, Tyler Buckingham

"Gold recovery gains in the order of one to two percent have been realized in 2009 due to changes in the operation of the grinding circuits at Barrick Gold Corporation’s Hemlo - Williams Mill. The initiative known as “split milling” involves milling low grade open pit ore in one grinding circuit and higher grade underground ore in the other grinding circuit under different conditions.The philosophy of split milling involves processing higher grade underground ore at a lower throughput rate and a finer grind while milling lower grade open pit ore at a higher throughput rate and a coarser grind. Bench scale tests showed that the recovery relationship for underground ore sources was much more grind sensitive than that of open pit ore. This relationship proved to hold in the plant. The gold recovery gain due to finer grinding of the higher grade ore is much greater than the recovery loss due to coarser grinding on the lower grade ore. Since the majority of gold is being recovered at a higher rate, this results in higher overall plant gold recovery when compared to the traditional operation philosophy of blending the ore sources.The challenges for achieving the maximum benefits of split milling were two fold; the need to increase throughput on the open pit ore milling circuit and a means of grinding finer on the underground circuit. The former was met with an increase in SAG (Semi-Autogenous Grinding) mill ball charge, a basic “drill to mill” blasting campaign and changes to the primary gyratory crusher. The latter was met with a reduced SAG mill ball charge and a higher milling density. The resulting benefits in additional gold recovered are in the order of $3-5 million per annum.INTRODUCTIONThe Williams Mill processes ore from three different ore sources, the David Bell Mine (DB), the Williams Underground Mine (WUG) and the Williams Open Pit (OP). All three ore sources are located within the Hemlo camp in Northwestern Ontario adjacent to the Trans-Canada Highway. The Williams Mill was built in the mid 1980’s at which time it was designed to process 3,000 dry metric tonnes per day through a 22’ SAG mill in series with a 16’ ball mill. In 1988 a Mill expansion doubled the mill capacity by adding a second identical grinding circuit. A photo of the current grinding circuit can be seen in Figure 1. Today the Williams Mill processes an average of 8,000 tonnes per day of which approximately 60% is from the Williams Open Pit."

Jan 1, 2010

By Mohamed Ourriban, Stéphanie Somot, Éric Proulx, Jean-Sébastien Marois, Marjolaine Drouin, Paul Blatter, Ahmed Bouajila, Claude Gagnon

"The mining industry has to face increasing restrictions concerning the use of fresh water in Québec and throughout the world. Increasing recirculated water usage could lead to a better control of the quality and the quantity of water supplied to the mineral processing units. Furthermore, it could reduce the volume of waste water to be managed and treated before its release to the environment. However, process water may have a negative impact on mineral processing operations, thereby limiting the incentive for sustainable development initiatives. As a first step into selecting and testing remediation strategies, one of which being blending local waters, a mapping of process water quality (inorganic and organic fractions) allowed to locate zones in an industrial case study where fresh water is introduced into the circuit and to detect dissolution and precipitation zones. A complementary hydro-geochemical modeling study on water blends was also performed showing encouraging results towards the objectives of the project.INTRODUCTION The industry has to face increasing restrictions on the use of fresh water in Québec (Quebec Government, 2009) and throughout the world (Brown, 2003; UNESCO, 2006). The cost for the use of fresh water and for water management is highly variable and case-dependent (from 0.2 to up to 7 million dollars/year according to available private estimations, excluding the capital expenditure). It may be, or may become, significant for any plant; thus, risk assessment is of concern here. Using recirculated water could lead to a better control of the quality and the quantity of water supplied to the mineral processing units. Furthermore, it could reduce the volume of waste water to be managed and treated before its release to the environment. However, process water may have some positive and negative impacts on mineral processing operations (Janssen, 2003; Johnson, 2003; Rao and Finch, 1989). Negative impacts could limit the incentive for sustainable development initiatives. Water management strategies at mineral processing sites are supported both by a strategic plan and a water management plan. The strategic plan includes legal aspects, investment costs, potential benefits and various risk levels (Corder and Moran, 2006; Côte et al, 2006; Evans et al., 2006; MacIntosh and Merritt, 2003; McLellan et al., 2008; Merritt and MacIntosh, 2006; Robinson and Keddie, 2006; van Beers et al., 2007a; van Beers et al., 2007b; van Berkel, 2007). In practice, the strategic plan evaluates technical possibilities and costs of various scenarios, such as for example: bringing fresh water to the plant or recirculating waters, or using blends of both. Two important factors to consider in order to enhance the future production efficiency and to minimise risks are: the durability and the stability of the resource, both from quality and quantity points of view."

Jan 1, 2012

By D. Imeson, T. Grammatikopoulos, G. Pearse, M. Aghamirian, C. Mohns

"Spodumene mineral is a source of lithium for lithium ion batteries. Thus, the production of a spodumene concentrate with an approximate grade of 6% Li2O is targeted. Traditionally, spodumene is concentrated by froth flotation using fatty acid as the collector. Amine collector can also be used to float mica ahead of spodumene as required or feldspar from the spodumene tailings. Slimes have a negative effect on spodumene flotation and they are commonly separated by scrubbing and de-sliming prior to spodumene flotation. The main losses in spodumene flotation typically occur in cyclone overflow, scrubbing slimes, spodumene flotation tailings and mica concentrate. In recent years, dense media separation (DMS) is the main concentrating process in Australian spodumene operations. In these operations, DMS is efficient at generating coarse grain concentrates and rejecting large portions of silicate gangue, thereby directly reducing the milling cost. The ore should be crushed to 3/8” or coarser while further crushing is possibly necessary on the middling particles to scalp further spodumene. The DMS feed should be classified to remove fines (typically -0.5 mm particles). The final middlings and undersize fractions could then be processed by flotation. The main requirement in DMS operations is to have good spodumene liberation at relatively large grain size. This paper reviews the most significant DMS and flotation parameters in spodumene beneficiation. The treatments of different minerals in the spodumene ore are also addressed.INTRODUCTION A number of laboratory/pilot plant programs have been recently conducted at the SGS Lakefield site, sponsored by different clients. The main objective has been to develop a process flowsheet to generate a spodumene concentrate containing 6.0% or higher Li2O (~75% spodumene) suitable for producing lithium carbonate by a hydrometallurgical process. Depending on spodumene liberation size, dense media separation (DMS) can be an attractive option to produce a final high grade concentrate and to reduce the mass to be treated in the flotation circuit. In such a case, crushing would typically be performed in stages, with DMS separation to recover a spodumene concentrate and reject gangue minerals following each stage. DMS middlings and -0.5 mm fines can be processed by flotation to further recover spodumene. In the flotation process, coarse stage grinding minimizes the grinding cost and formation of slimes. In the majority of cases, a grind size (K80) of approximately 200 microns is sufficient for flotation. After primary grinding, mica, if present in an appreciable amount, should be floated to minimize interference with spodumene flotation. Desliming after scrubbing and high density conditioning are the key process operations prior to spodumene flotation. Spodumene can then be floated from silicate gangue minerals with a fatty acid or a mixture of fatty acid and petroleum sulfonate at slightly alkaline pH. Subsequently, feldspars can then be floated from the spodumene tailings."

Jan 1, 2012

By David Whalen

Mobile equipment in oil sand mining operations experience more maintenance problems than similar equipment in other surface mines, due to soft underfoot conditions. Oil sand properties are known to vary with temperature, bitumen content and previous loading cycles. One of the challenges is the widely variable underfoot pressure which supports shovels and trucks. In the case of a heavy hauler driving on in-pit roads, each tire is supported by a different ground stiffness resulting in an apparent uneven surface. Driving over uneven ground causes twist in the truck frame, which is transmitted to the dump body. Through a finite element analysis of a haul truck dump body, it was shown that slight frame twists produce localized stresses in the intersections of the floor stiffeners. These stresses are at times greater than those resulting from the maximum ore weight and are primarily responsible for cracking in these intersections. Syncrude maintains high quality haul roads, which are constructed of gravel, and makes every effort to keep the shovel pit roads as smooth as possible. This problem cannot be avoided through operational practices only and must also be accounted for in the design of haul truck bodies.

May 1, 2003

By L. Nichols, H. Keele, K. Czajewski

"The oil sands industry in northern Alberta disposes of the by-products of bitumen extraction into tailings ponds. The mix of tailings sand and mature fines, with gypsum added as a flocculent, constitute a waste stream called consolidated tailings or composite tails (CT). The two main challenges of the tailings disposal process are a chronic shortage of economical containment and the potential mobility of CT deposits in the event of breaching by, for example, long-term erosion, which makes abandonment of tailings ponds more complicated. Thick deposits of CT do not consolidate in adequate time to provide reassurance for abandonment of the CT containment facilities and that is why the containment must be adequately robust or the properties of CT improved for long-term stability. This paper addresses potential solutions to these problems by applying the gravity drainage of CT and quasi-homogeneous construction for containment facilities.Gravity drainage is suggested as a mechanism to facilitate the consolidation of CT in critical areas of the ponds by inserting internal sand layers within the body of the deposit. Critical CT areas are defined as areas around the perimeter of containment. The advantages of this solution are longterm stability of the disposal area by creation of stiff deposits around its perimeter and facilitation of reclamation efforts on its surface.The present solution for containment of CT deposits is not favourable from a materials balance point of view. In general, sand has become a valuable resource as it both forms the primary component of the CT matrix and becomes the containment building material. If sand is required for CT, less is available for dyke construction. This puts additional pressure on the industry as a whole to maximize the use of overburden and interburden for construction of tailings containment structures. Typically, only 50 percent of dry mine waste (overburden + interburden) is incorporated into dyke construction. The remainder of dry waste is stored either in waste dumps or in tailings ponds, consuming valuable in-pond space originally allocated to CT material. Quasi-homogeneous dykes are proposed to contain CT deposits rather than traditional minimum cross-section dykes built of overburden and/or cell sand. The quasi-homogeneous dyke may change the unfavourable material balance and lessen the construction constraints that are presently imposed on the mine/waste disposal process."

Jan 1, 2008

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

By Norman O. Lotter

Drill core is preferred as the sample material for metallurgical and mineralogical characterisation and flowsheeting of an orebody. The question arises as to how to do this at the economic optimum, since geologists have several other strategic uses for their drill core. The limiting factors are the cost of extra drilling for the metallurgical testwork and the need to preserve the drill core record. In this study, the representative sampling of drill core at an economic optimum for metallurgical testing is addressed. This phase of the drill core work is preceded by earlier drill core characterisation - for which other appropriate sampling protocols are used - that has provisionally categorised the orebody into geometallurgical units. In stratified sampling of drill core, use is made of a version of the stratified approach to compound distributions proposed by Cochran in 1946. Use is made of Gy’s 50-piece minimum sample mass model to estimate minimum requisite sample mass. In the case where sampling and testwork are required on a Lifeof- Mine blend of a given ore deposit, whereby defined ore types or geometallurgical units are present, the sample mass requirement is significantly reduced by working from drill-core at the geometallurgical unit level. A case study using Raglan drill core is discussed. Using nickel as an indicator paymetal, and sampling at a fundamental variance level of 8%, the minimum sample masses for disseminated sulphides, net-textured sulphides, and massive sulphides are 3.80, 7.37 and 4.39 kg respectively. When the same drill-core is sampled as a Life-of-Mine mixture without knowledge of the geometallurgical units, the minimum sample mass increases to 84.28 kg. It is thus possible to assemble a representative Life-of-Mine sample of drill-core more economically, at the same time validating its representativity. Equivalent calculations for copper, platinum and palladium are discussed, with outcomes consistent with this finding. When all four paymetals are considered, one selects the largest minimum mass out of all four cases. This results in minimum sample masses of 11.67, 74.90 and 175.16 kg for disseminated sulphides, net-textured sulphides and massive sulphides respectively. The simulated Life-of- Mine blend would require a minimum sample mass of 377.33 kg to cover all four paymetals.

Jan 1, 2010

By Joey Hoang, Robert Johansson, Tapani Keronen, Mikael Jåfs, Jani Jansson, Mikko Korpi

At the Copper 2016 conference in Kobe, Japan, Outotec introduced a vision for the year 2020, wherein smelting would make use of four digital cornerstones: intelligent instruments, advanced process control, improved information management and transparency of and access to data. Since then, this vision coined the phrase “Fully Automated Smelter” with the mission promising a concept where no manual actions from human operators would be needed during normal smelter operations. Replacing repetitive physical work, collection and analysis of everyday production data has a great potential for further automation and will provide productive ways to improve efficiency and safety of smelters in the future. Therefore, a lot of emphasis has been put into developing solutions to automate these tasks and to move people out of harm’s way. A lot of new equipment, systems and optimization tools are under development. In addition, a lot of effort is going into developing dynamic process models and advanced control systems to interface with the plant automation systems. The goal is to integrate all these concepts to meet Outotec’s Fully Automated Smelter vision. This paper will cover these latest achievements toward this vision and introduce the near-future of digital smelting. INTRODUCTION As is well known already, the humankind is currently experiencing a change, extent of which is broadly studied by the governmental, educational and industrial institutions and parties. The wellbeing and quality of life have traditionally been improved alongside the developments in industrial production, which is now about to enter its fourth revolution. Therefore, the current change is often described as fourth industrial revolution, or industry 4.0 (or industrie 4.0, as named by Acatech, 2013) and can be considered anything from small incremental development to a ground-breaking global reset that will affect everyone’s lives. Many seem to agree, that the scale of change will indeed be big, even significantly bigger than the three industrial revolutions before it (Pellini, Weyrauch, Malho & Carden, 2019; Geissbauer, Vedso & Schrauf, 2016; Jansson, Bjorklund, Lahtinen, Heinonen, Jafs & Fagerlund, 2016).

Jan 1, 2019

By J. E. Nesset

"Significant progress has been made over the past few years on characterizing and understanding the hydrodynamic properties of frothers and collectors and their effect on flotation performance. This new understanding of the relationship between chemistry and hydrodynamics has dramatic implications for how flotation circuits should be controlled and optimized. Previous work has shown that choosing a frother of the right ‘strength’ is critical to the optimization process, and the work and discussion presented here serve to strengthen the argument. In addition, we know that collector chemistry will also influence hydrodynamic properties and needs to be taken into consideration. To achieve the desired outcome, the overall system of reagents and operating parameters must be considered. This paper describes and discusses some of the findings from recent research on frother and collector chemistry effects (on hydrodynamics), and how this knowledge can be used to improve flotation performance. Examples from plant experience are discussed.INTRODUCTIONFrothers perform various functions in flotation but the primary ones are control of bubble size in the pulp zone and stabilization of the froth zone. Laskowski (2003) was among the first to try to capture both roles in characterizing frothers in order to provide a basis for frother selection for a given duty, as well as to explore the link between function and frother chemistry (structure). The method he used requires bubble size data and a frothing index; the procedure has been simplified by Cappuccitti and Finch (2008) using gas hold-up to substitute for bubble size, and froth height (without overflow) as the froth stability metric. The gas hold-up/froth height relationship will be referred to as ‘hydrodynamic characterization’. Essentially, the gas hold-up parameter reflects the sum of the hydrodynamic conditions in the pulp phase (water in this case), i.e. the effect of bubble size plus the apparent impact of the water surrounding the bubble due to the frother (Finch, Nesset and Acuna, 2008). Similarly, the stable froth height parameter reflects the sum of hydrodynamic conditions in the froth, i.e. the effect of bubble size undergoing coalescence plus the influence of the water in the lamellae and Plateau borders between bubbles. It could be argued that the gas hold-up/froth height relationship is a superior one to simple bubble size/froth index for discerning differences in frother behaviour"

Jan 1, 2010