Search Documents

By M. Abzalov

"Control of analytical data quality is usually referred to in the mining industry as Quality Assurance and Quality Control (QAQC), and involves the monitoring of sample quality and quantification of analytical accuracy and precision. QAQC procedures normally involve using sample duplicates and specially prepared standards whose grade is known. Numerous case studies indicate that reliable control of sample precision is achieved by using approximately 5% to 10% of field duplicates and 3% to 5% of pulp duplicates. These duplicate samples should be prepared and analyzed in the primary laboratory.Bias in the analytical results can be identified by inclusion of 3% to 5% of the standard in each sample batch. Several different standards are used, with values spanning the practical range of grades in the actual samples. A blank (a sample in which the concentration of metal of interest is below detection limit) should also be included. Standard samples alone cannot identify biases introduced during sample preparation, and therefore approximately 5% of the duplicate samples (coarse rejects and pulp) should be processed and assayed at another, external, reputable laboratory.This paper discusses techniques used for estimation of errors in precision and accuracy, and overviews diagnostic tools. It is shown that one of the most commonly used methods, the Thompson-Howarth technique, produces consistently lower results than other methods. These results reflect the nature of this method, which relies on the assumption of a normally distributed error, and thus produces biased results when errors have a skewed distribution. This study concurs with the suggestion of Stanley and Lawie (2007: Exploration and Mining Geology, v. 16, p. 265-274) to use the average coefficient of variation (CVAVR(%)) as the universal measure of relative precision error in mine geology applications: Le contrôle de la qualité des données analytiques est habituellement désigné sous l’appellation d’Assurance et Contrôle de la Qualité (ACQ), elle implique le suivi de la qualité des échan¬tillons et la quantification de la reproductibilité et la précision analytiques. Les procédures des pro¬grammes d’ACQ impliquent habituellement l’usage de duplicata et de standards analytiques dont la teneur est connue. Plusieurs études de cas indiquent qu’une bonne fiabilité du contrôle de la précision des échantillons est atteinte par l’usage de 5% à 10% de doublons sur le terrain et de 3% à 5% de doublons de pulpe. Ces duplicata devraient être préparés et analysés au laboratoire principal.Un biais des résultats analytiques peut être identifié par l’inclusion de 3% à 5% d’échantillons standard dans chaque envoi d’échantillons. Plusieurs standards sont utilisés, avec des valeurs couvrant l’étendue des teneurs habituellement rencontrées dans des échantillons réels. Un blanc (un échantillon dans lequel la concentration d’un métal donné est sous la limite de détection) devrait aussi être inclus. L’emploi exclusif d’échantillons standards ne permet pas l’identification de biais résultant de la préparation des échantillons, de telle sorte qu’environ 5% des duplicata (rejets grossiers et pulpes) devrait être traités a un autre laboratoire indépendant et fiable.Cet article passe en revue les techniques utilisées pour évaluer l’erreur liée à la précision et a la reproductibilité, et effectue aussi un survol de divers outils diagnostiques. Il ressort de cet examen qu’une des techniques les plus fréquemment utilisées, la méthode Thompson-Howarth, fournit des estimés systémati¬quement inférieurs à ceux provenant des autres méthodes. Ces résultats sont un reflet de la nature de cette méthode, laquelle assume une distribution normale de l’erreur, biaisant ainsi les résultats lorsque l’erreur présente une distribution asymétrique. La présente étude recommande de privilégier l’usage du coefficient moyen de variation (CVAVR(%)) en tant que mesure universelle de la précision relative de l’erreu"

Jan 1, 2008

By D. M. Fraser

"Brunswick Mining and Smelting Corporation Limited operates two mines; one open pit (No. 6) and one underground (No. 12), approximately 20 miles south-west of Bathurst, New Brunswick. Both are massive sulphide deposits containing copper, lead, zinc and silver. The No. 12 property began operation in 1963 and is presently being mined at a rate of 3600 tons per day underground and is in the completion stages of an expansion to achieve 11,000 tons per day underground by the end of 1930.Originally at Brunswick Mining, acid mine drainage was added to tailings and neutralized by an excess of lime in the tailings. A conventional recycle system was established with a biostabilization pond for the treatment of thiosalts, a liming station and sludge settling area for neutralization of acidic water and a final reclaim pond with a pumping station for reclaim water.We found that not more than one third of the reclaim water could be used as it upset the flotation circuit every time it was used. It was established that the mine water and surface drainage were the cause due to oils and detergents in the water. It was also found that the biostabilization pond method Cf thiosalt treatment was not efficient for year round operation in our climate.In 1974, we reviewed the problems and began a 3 phase program to control effluents from the property:"

Jan 1, 1980

By G. M. Ritcey, K. E. Haque, S. H. Lucas

"Laboratory-scale hydrochloric acid leach tests were conducted on an Elliot Lake uranium ore. Hydrochloric acid leaches without an oxidant or with molecular oxygen (i.e., air or pure oxygen gas) as an oxidant favoured uranium and radium extractions. Conditions for maximum uranium, thorium and radium extractions with hydrochloric acid in the presence of sodium chlorate were determined. The highest extractions of uranium (96%), thorium (81%) and radium (91%) were obtained using a mixture of 44.0 kg HCI and 2.5 kg NaCIOJ per tonne of ore at 75°C with 18 hours ' retention.Flowsheets for two-stage leaching are proposed. The residues obtained from the second-stage leach contain uranium (0.002%), thorium (0.006%) and radium (15 to 20 pCi/g). Tailings of this nature are considered suitable for mine backfill, and, if pyrite is removed, suitable for surfacedisposal.IntroductionThe nuclear energy industry forecasts an increasing demand for uranium. This increasing demand requires not only largescale mining and processing of conventional and complex uranium ores, but also leads to the consideration of new and nonconventional sources of uranium, e.g., black shales, phosphate rocks, granite, sea water and fossil-fuel ash.Sulphuric acid leaching is the most widely used process to extract uranium from the ore(l). The leach liquor is separated from the solids (tailings), purified, concentrated in various intermediate stages, and finally a semi-purified product, yellow cake (80-85070 U30 8) , is precipitated. The tailings, which constitute nearly 98% of the total bulk of most ores, are usually discarded in surface disposal areas.The Elliot Lake region of Ontario is one of the major uranium producing areas in Canada. The principal uranium minerals are brannerite and uraninite, present as microscopic grains in the matrix of quartz-pebble conglomerates. The average grade is between 0.10 and 0.15% UPs. Up to 6.5% pyrite is also present in the conglomerates. Because pyrite is less reactive toward sulphuric acid leaching, almost all the pyrite goes into the tailings and eventually into' the disposal sites."

Jan 1, 1980

By Manqiu Xu, Peter Wells, Gord Bacon

"A methodology is proposed for milling flowsheet design using flotation rate, selectivity and residence time distribution data. The flotation rate data and the difference in flotation rates between the main minerals to be separated are determined from laboratory batch tests. The residence time distribution (RTD) data is calculated using the tanks-in-series model assuming each flotation cell is a perfect reactor. The rate, selectivity and RTD data can be used to calculate the optimum flotation cell number in a bank for rougher/scavenger in a commercial plant.This methodology is demonstrated in a case study of a nickel sulfide ore. It is shown that miniplant rougher/scavenger data can be closely predicted from a batch test using the proposed methodology.INTRODUCTIONIt has been the trend in the past several decades that large flotation cells are increasingly used. This is largely driven by the economics as ore grades are getting lower and the rate of ore processing must be increased in order to keep mining operations economically viable. The development and evaluation of 250 to 500 m3 WEMCO SmartCell and Outokumpu TankCell flotation machines are well in progress. There are many advantages in using large cells: fewer cells, less floor space, lower power cost (although this is now being questioned for fine particle flotation), simpler circuit, simpler control, and lower auxiliary equipment cost (Arbiter, 1999). As flotation cells are getting larger, the importance of cell hydrodynamics, namely, air/slurry dispersion, power intensity, froth transfer, residence time distribution (RTD) and short-circuiting are being fully recognized (Jonaitis, 1999; Weber, et al., 1999; Gomez and Finch, 2002). It is not the intention of this paper to discuss flotation cell scale-up. The intention is to answer what is the optimum number and size of cells in a flotation bank for a given throughput. Figure 1 illustrates this problem facing design engineers. The same mean residence time required for a given throughput can be achieved in a single large cell, or many smaller cells in a row. A few cells of large size may be more cost-effective but they may not be as effective as more cells of smaller size in achieving the optimum metallurgy."

Jan 1, 2005

By M. du Plessis, G. More O’Ferrall, A. Hartzenberg

"Lonmin Platinum has experienced several cases of large-scale hangingwall instabilities during the extraction of the Upper Group 2 Chromitite Layer (UG2) at a number of the Marikana operations along the Western Limb of the Bushveld Complex. These instabilities were all attributed to the presence and interaction of prominent geological features and highlighted that assumptions made in the design of the stable pillars and stable span were potentially overly simplistic in these areas. Investigations into the various case studies over the past ten years have resulted in the identification and classification of the rock mass characteristics of the geological structures causing the instabilities.In order to get a better understanding of these instabilities, it is important to understand the structural geology in general, and the sequence of events that led to the formation of the Bushveld Complex in particular. The tectonics involved in the formation of the various layers around the economic mineralized layers has been investigated, and the influence of these tectonic events on the stability of the UG2 stopes hypothesized.Following this investigation, the behaviour of the various layers in the hangingwall of the UG2 is better understood. This paper presents a description of the ground conditions being experienced in the underground stoping environment, the possible source of the associated structures and the resultant behavioural mechanism of these structures causing the rock mass instabilities."

Jan 1, 2015

By P. Kolodziejczyk

Mining seismicity can be harmful both under and on the surface. ? Tremors are not totally damaging, but destructive to buildings and infrastructure, especially in highly urbanized regions, such as Upper Silesia in Poland ? Destroy the social appreciation, trust and support for mining industry ? To prevent or reduce these effects, engineers and scientists develop methods allowing prediction of ground shaking intensity near buildings and various losses that can result from it Evaluation and prediction of mining tremor?s damaging influence on buildings is usually solved with two steps: ? Estimation of local ground shaking intensity (based on known attenuation relation) and classification into one of several predefined intensity intervals ? Apply supposed relation between the shaking intensity interval and qualitatively defined damaging effects to express predicted effects This makes a mixed quantitative/qualitative procedure, difficult to analyze or optimize

Aug 1, 2013

By S. Hoskins, S. Chiovelli

"The annual budget for the repair and maintenance of equipment at Syncrude Canada Ltd. is in excess of $450 million. Of this amount, a significant portion can be attributed to the loss of material on various pieces of equipment throughout the operation. In order to effectively battle the different wear and corrosion mechanisms that lead to material loss, it is important to understand how materials behave in the environments in which they are used. Through the work carried out at Syncrude, and work conducted with various external collaborators, significant improvements in both equipment reliability and productivity have been achieved. This paper will review some recent field trials where materials research has demonstrated the potential for significant improvements in equipment run life and reduced maintenance costs. These field trials will demonstrate the need for understanding the interaction between various damage mechanisms and the materials used in different services. The paper will also discuss current areas of focus along with future directions for further work. IntroductionSyncrude Canada Ltd, located in northern Alberta, Canada, operates a mining, extraction, and upgrading facility that removes bitumen from oil sand to produce synthetic crude oil. Oil sand is mined from the orebody using conventional open pit mining techniques. Recently, the mining scheme has changed from a dragline/bucketwheel/conveyor system to a shovel/truck/crusher/hydrotransport operation as depicted in Figure 1. Since the start of production in 1978, the identification, development and application of wear protection has been an ongoing challenge due to the abrasive nature of the oil sands and the massive volumes of materials that are moved. The use of various materials and application processes along with component design modifications has resulted in significant improvements in both productivity and reliability (Adamic et al., 1995; Llewellyn, 1996, 1997; Anderson et al., 1998; Chiovelli et al., 1998).In an effort to continuously improve the operation, Syncrude has been quick to adopt new wear technologies. As seen in Figure 2, since the start of operation, a variety of materials have been used to increase wear performance and ultimately equipment reliability and availability. Although Figure 2 does not show a time axis, the bottom axis may also be used to represent a time progression for the predominate use of these materials at Syncrude. Tungsten carbide metal matrix composite (WC-MMC) overlays are currently the materials of choice for most of Syncrude’s critical production applications. Metal matrix composite overlay technology has potential for significant improvements and Syncrude is actively working with powder consumable manufacturers to advance this technology."

Jan 1, 2004

By Michael Young, Josh Rubenstein, Michael Cooper, Tom Shouldice, Dan Curry

"Xstrata Nickel Australasia’s Cosmos plant in Western Australia is currently implementing a debottlenecking project to increase nickel output while maintaining metallurgical performance and a low unit cost for production. Mineralogical analyses indicated that the processing bottlenecks could be removed by simply reconfiguring portions of the circuit as well as adding capacity with technical advancements – notably an IsaMill™, Jameson Cell, Wemco® SmartCell™ and Larox Filter.The Wemco® SmartCell™, which is particularly suited for coarse mineral flotation, and Jameson Cell, which is particularly well suited for fast floating and fine minerals, combine as rougher scalpers to complete the main stream portion of the flotation circuit. Addition of the two new cells meet the required flotation capacity to handle the forecast increase in nickel feed grade and tonnage.The addition of an IsaMill™ for regrinding allows the SAG mill to produce a coarser primary product, thus enabling increased throughput. The advantages offered by the IsaMill™ include operation in open circuit (internal classification), an inert grinding environment (ceramic media) for better downstream flotation (“clean” new surfaces), compact footprint (easy to retrofit into existing plant) and energy efficiency. With the new rougher capacity addition, the remaining original plant cells were reconfigured to suit the new regrind mill and cleaning requirements. The existing 24 m2 Larox pressure filter was replaced with a larger 32 m2 unit to dewater the increased concentrate output.The paper describes the upgrade at the Cosmos concentrator with particular emphasis on the mineralogical data and metallurgical benefits associated with the changes.INTRODUCTIONThe Cosmos concentrator is located 680 km northeast of Perth in Western Australia. The deposit’s development and operations were started in 1999 by Jubilee Mines NL, Cosmos’ previous owner, with the first production of nickel concentrate in April 2000. The Cosmos concentrator treats a komatiite hosted high grade nickel sulphide ore, with sulphides being predominantly pentlandite and pyrrhotite with lesser amounts of pyrite and chalcopyrite."

Jan 1, 2010

By R. Simonson

Agenda ? Need and Introduction ? Interface Management Programs ? Outcomes of Successful Interface Management ? Key Success Factors ? Lessons Learned ? Concluding Remarks Need for Interface Management ? Complex projects ? Larger in scale ? Physically dispersed and delivered remotely ? Involved with several stakeholders ? Risks thus imposed due to: ? Iterative and dynamic nature ? Large information flow ? Document revisions and information updates ? Related sources of Risk: ? Ineffective communication ? Poor work relationships ? Person to job mismatches ? Failure to clarify expectations

Aug 1, 2013

By Edgar Stanfield

Since April, 1923, the Scientific & Industrial Research Council of Alberta has had an analyst, Mr. W. P. Campbell, steadily employed analyzing coal samples taken by the provincial Mine Inspectors. Mr. Campbell has also analyzed a number of coal samples taken by one of the Council's Geological Survey parties. We have thus, at the time of writing, over four hundred analyses of Alberta coal, made by the same analyst, and strictly comparable. These samples come from widely distributed points, some 88 townships being represented. The stream of samples is still flowing in, and it is hoped by the end of this year that the number of townships represented will be increased at least fifty percent. The time has arrived to publish some of the data collected; but it must be clearly understood that this is merely a progress report, subject to revision as further information is obtained. Three hundred and fifty-three coal mines in the province were operated during 1923. The production per mine varied from 5 up to 500,000 tons. With this number of mines to examine (and fresh mines are being opened every year) it is a large task to make even a proximate analysis of representative samples from every mine. Any chemist will appreciate the fact that to make ultimate analyses representative of even the larger mines will be a long and tedious task. It has seemed better to make a comprehensive survey by proximate analysis and determination of calorific value before beginning the more thorough survey by ultimate analysis. This paper, therefore, deals only with proximate analyses and calorific values; and, owing to limitations of time and space, does not deal with other chemical and physical characteristics of coal, such as storage quality and fusibility of ash, important though these are.

Jan 1, 1925

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 Zhibao Li, Kunming Zhai, Yan Zhang

The hydration kinetics of MgO, namely magnesia or magnesium oxide, obtained by calcination of magnesite MgCO3 or synthetic MgCO3·3H2O has been studied. Nesquehonite (MgCO3·3H2O) was prepared by the reaction of MgCl2 with Na2CO3 in a mixed-suspension mixed-product removal (MSMPR) crystallizer at 40 °C. Three magnesium oxide samples were obtained by calcination of magnesite MgCO3 or synthetic MgCO3·3H2O in muffle furnace at 900, 1100 and 1300 °C for 3 h, respectively. The hydration experiments of MgO samples were investigated at 25, 35 and 45 °C for 24 h in a stirred batch reactor. The hydration results indicate that the rate and extent of hydration of magnesia decreased remarkably with hydration temperature and increased with decreasing calcination temperature at a constant hydration temperature. The whole hydration reaction of magnesia was supposed to experience the promotive period and declining period. An empirical kinetic model ( k t p -?ln(1-??) ??) and a modification of the classical shrinking core model for diffusion control ( k t b d -?-?3 2 ????1 [1 (1 ??) ] ) have been applied to fit the experimental data of hydration in two stages, respectively. The values of kinetic parameters ( p k and d k ) were obtained by regression analysis and the activation energy of the sample at 900 °C was estimated to be 34.23 kJ/mol. The comparison between MgO from calcination of synthetic MgCO3·3H2O and a commercial MgO used in silicon steel production was carried out, indicating that special silicon steel grade MgO can be obtained by calcination of MgCO3·3H2O below 900 °C.

Jan 1, 2014

By V. Lakshmanan, Dan Mackie, J. Melnbardis

"Crystalline Flake Graphite is a strategic mineral. Since 1954 nearly all North American consumption has been imported from offshore. The Bouthillier Project in Quebec, Canada, is Orrwell Energy's bid to enter the market, with plans to begin production of a 500 TPD operation in 1984.Milling of flake graphite is considered easy, yet more problems have been encountered in flowsheet and plant design than in most other minerals. A fresh approach to existing technology was taken, giving results that are highly satisfactory. The key to graphite liberation without flake deterioration is found to be in autogeneous grinding. In addition to producing marketable coarse graphite by such autogeneous grinding, flotation and tabling, super purity fines can be produced using repeated regrinding, elutriation and flotation. This paper discusses the design of a 500 TPD plant based upon the results obtained from bench and pilot plant studies. INTROOUCTIONWhen the Black Oonald mine at Calabogie Ontario ceased production in 1954 nearly all North American crystalline flake graphite had to be imported from offshore. Orebodies in the United States were too small or too low in grade to be exploited economically.The Houthillier Township near Mont Laurier in Quebec (Figure 1) was a known graphite-bearing area The Orrwell property in that area was delineated in 1957 Some test work including minor footage of diamond drilling was carried out by Italia Copper and the MHM Syndicate prior to 1981 when Orrwell acquired the property"

Jan 1, 1984

By Zetti Gavelán

Lack of skilled workforce in mining industry during the current decade has been a serious concern for most mining organization globally. At the beginning of this decade the major envisioned difficulty was related to generational breach of mining professionals. By 2003, mining industry recover from a low metal prices cycle following by a worldwide mining investment boom in 2005 that triggers a talent war among mining companies that was not only limited to professional level but skilled operators as well. Despite world?s economy collapse in October 2008, impact on mining industry was small comparing to other industries and appears to show evidence of healing. Under this context, fundamentals that have been driven lack of skilled workforce for the mining industry in previous years still valid. Starting in 2007, Minera Argentina Gold SA - Barrick Veladero, subsidiary of Barrick Gold, initiated an innovative selection and training program for candidates without exposure to mining. Following our corporate social responsibility mandates, Barrick Veladero decided to change its hiring process entirely, concentrating on selection and training of local people with no experience rather than hire experienced mining operators. This paper shares Barrick Veladero experiences on its hiring program, challenges faced, and current results.

May 1, 2010

By G. M. Ritcey, K. E. Haque, S. H. Lucas

"Laboratory-scale hydrochloric acid leach tests were conducted on an Elliot Lake uranium ore. Hydrochloric acid leaches without an oxidant or with molecular oxygen (i.e., air or pure oxygen gas) as an oxidant favoured uranium and radium extractions. Conditions for maximum uranium, thorium and radium extractions with hydrochloric acid in the presence of sodium chlorate were determined. The highest extractions of uranium (96%), thorium (81%) and radium (91%) were obtained using a mixture of 44.0 kg HCI and 2.5 kg NaCIOJ per tonne of ore at 75°C with 18 hours ' retention.Flowsheets for two-stage leaching are proposed. The residues obtained from the second-stage leach contain uranium (0.002%), thorium (0.006%) and radium (15 to 20 pCi/g). Tailings of this nature are considered suitable for mine backfill, and, if pyrite is removed, suitable for surface disposal.IntroductionThe nuclear energy industry forecasts an increasing demand for uranium. This increasing demand requires not only largescale mining and processing of conventional and complex uranium ores, but also leads to the consideration of new and nonconventional sources of uranium, e.g., black shales, phosphate rocks, granite, sea water and fossil-fuel ash."

Jan 1, 1980

By Bern Klein

Motivation for Improved Energy Efficiency in Mining ? Canada potentially adopting Greenhouse Gas Emissions and Intensity Legislation in 2013 ? BC -Climate Action Plan and Energy Conservation Agreement in 2008 ? BC Energy Plan 2007 -BC Hydro is mandated to save 50% of new load growth ? TSM Initiative -Energy Management, Energy Reporting, Setting Energy Targets, GHG Management BC MOU for Energy Conservation Agreement -MEMPR, MABC, BC Hydro

May 1, 2010

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 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

By F. Bruey, T. Bruce, D. R. Nagaraj, W. Thomas, P. Riccio, E. Guitard

"AEROPHINE® 3418A Promoter is a flotation collector that has been an industry standard in the Canadian mineral processing industry since its first commercial use at the Kidd Creek concentrator in 1980. Its wide range of applications in sulphide mineral flotation has made it successful, specifically in the processing of complex sulphide ores which contain copper, lead, zinc, and silver or any combination of these elements for selective separation.AEROPHINE® 3418A Promoter is a unique, phosphine-based collector originally developed for the flotation of copper sulphide minerals and activated zinc minerals. It is now recognized for its selectivity for complex, polymetallic sulphide ores and porphyry copper ores. In addition, AEROPHINE® 3418A Promoter is highly selective for copper, lead, activated zinc, nickel and precious metals. AEROPHINE® 3418A Promoter exhibits strength yet, selectivity against iron sulphides (pyrite, pyrrhotite, marcasite, and arsenopyrite), non-activated sphalerite, and many minerals which contain penalty elements.This paper outlines a brief history of the product’s development, its historical use in Canada and several potential innovative mineral processing applications with the use of the chemistry. Three Canadian case studies will be presented on various ore types including: a Cu-Au ore, a Cu-Zn ore, and Cu-Pb-Zn ore. A plant production update will also be given on the future supply of AEROPHINE® 3418A Promoter.INTRODUCTIONCytec Industries Inc. operates a phosphine derivatives plant in Niagara Falls, Ontario, that manufactures various products for their speciality chemicals division. The products from this site are based on phosphine (PH3) chemistry and are used in a variety of markets including electronics, agriculture, LED lighting, water treatment, metal extraction, and mineral processing. One of the major products is the phosphine-based sulphide mineral collector, AEROPHINE® 3418A Promoter.AEROPHINE® 3418A Promoter (Figure 1) was originally developed for the flotation of copper and activated zinc minerals (Thomas, 2010).."

Jan 1, 2015

By Nick D. Rose

Outline ?Background ?History of inverse-velocity for failure prediction ?Methodolog ?Case Histories ?Three case examples of predicted slopefailure to actual date of occurrence ?One case example of mitigated slope failure using inverse-velocity to forecast failure time and schedule remedial measures ?Conclusions

May 1, 2006