Search Documents

By H. E. Urtubia, C. W. Bailey

"BHP-Billiton’s Minera Escondida Limitada is one of the world’s largest copper producers. It treats a combined 88 million tons of copper ore per year through its two concentrators: Los Colorados and Laguna Seca. Over the past three years, metallurgical sampling has been performed at the Laguna Seca concentrator to calibrate both the JKSimMet and JKSimFloat grinding and flotation simulators. In the past, the calibrated simulators have been used to predict the effect of changes in operating conditions in the grinding and flotation circuits separately. To assist in the development of future processing options, predictions were required of the effect of changing grinding circuit conditions on flotation performance. This was achieved by using an integrated grinding and flotation simulation. The grinding circuit performance was predicted under different operating conditions (such as tonnage rate and product grind size) using JKSimMet, and these results were then used in JKSimFloat to predict the flotation performance for the new grinding circuit product. This combination of the simulators allows both the effects of residence time and grind size on flotation performance to be investigated.This paper presents the first combined grinding and simulation modelling and simulation case study using JKSimMet and JKSimFloat. The modelling and simulation methodologies for the grinding and flotation circuits are discussed along with the circuit simulations that were performed. The simulator predictions are presented in conjunction with independent laboratory results of the new operating conditions. The ability to predict flotation circuit performance based on changes in grinding circuit performance is also discussed. This methodology provides improved predictions of flotation circuit performance when grinding circuit changes (including grinding capacity increases) are made.INTRODUCTIONOver the past decade, the grinding and flotation research groups at the Julius Krutschnitt Minerals Research Centre - JKMRC (Brisbane, Australia), University of Cape Town (Cape Town, South Africa) and McGill University (Montreal, Canada) have been developing a methodology for optimising grinding and flotation plants using actual plant data together with semi-empirical sub-process models. The project is being conducted as part of the Australian Minerals Industry Research Association (AMIRA) P9 Project, titled “The Optimisation of Mineral Processes by Modelling and Simulation”. The project first commenced in 1962 and is currently in its “Nth” phase."

Jan 1, 2007

By Erin N. Legault-Seguin

The Lac des Iles Mine (LDI) is located approximately 100 km northwest of Thunder Bay, ON. The mine has been in operation since 1993. The initial mine production was 2,300 mtpd and increased to 15,000 mtpd following a mine expansion project that was completed in 2001. Effective Oct. 29, 2008, LDI was shut-down under care and maintenance due to low metal prices. Prior to shutting down, LDI was the only primary PGM deposit in operation in Canada and of the few found globally. The complex geology and physical properties of the deposit causes several challenges for the concentrator. This paper will discuss the solutions to the main challenges of attaining nameplate tonnage and achieve satisfactory Pd recovery and concentrate grades.

May 1, 2009

By L. M. Pidgeon

MAGNESIUM metal, with a specific gravity of 1.73, is two-thirds the weight of aluminium and one-quarter that of steel. It is the lightest metal which is stable in the atmosphere. Before the war it had begun, though somewhat slowly, to take a place among the established metals. In 1939, magnesium production on this continent amounted to some 3,000 tons per annum and was confined to one company. At that time, Germany produced perhaps 20,000 tons. The advent of war, however, increased the demand to fantastic heights and introduced production problems of unprecedented magnitude. By the end of 1943, North American production is expected to reach 300,000 tons per annum (1), an increase of one hundredfold. Such a rate of expansion has not been witnessed previously in the history of metallurgy. It has given the metallurgist an opportunity to try new production methods and to learn in a short time what might otherwise have taken many years. The metallurgy of magnesium is still in its formative years and at this time widely different production methods are in operation side by side. This paper will discuss these methods in so far as information is available for publication at this time. A general description of the new Canadian industry will be included.

Jan 1, 1944

By Linda Duncan, Trevor Yeomans, Chris Potorieko, Michael Fullam

"NVI Mining’s Myra Falls Operation is a 4,000 mtpd copper/zinc flotation circuit with by product gold of ~28,000 ounces per year. Gold recovery has historically averaged 40-55%, however, a portion of this was recovered in the zinc concentrate and received poor payment. Two Knelson CD-30 Concentrators have been operating in the grinding circuit since 1992, with gravity recovery averaging about 2-5% over that time period.The Myra Falls gravity circuit was one of the very first hard rock mining installations operating in the world with automated Knelsons. The mine has employed a unique and unusual upgrading method for the Knelson concentrate that had never been evaluated as to its efficiency. Recent Gravity Recoverable Gold Testing by McGill University, and mathematical modelling of the Gravity Circuit by Knelson Gravity Solutions indicated the gravity circuit was not living up to its potential. Myra Falls Personnel changed the upgrade system-operating regime, and a complete gravity audit of the grinding circuit was carried out, with the results confirming the initial suspicions. Gravity recovery increased from 5% to 15%, with overall gold recovery improving by 5%.INTRODUCTIONNVI Mining operates a 4000 tpd mine and mill complex on Vancouver Island. The milling circuit consists of open circuit rod milling, closed circuit ball milling, gravity recovery, copper flotation, zinc flotation, and tails disposal. There are two milling circuits, “A” circuit, and “B” circuit."

Jan 1, 2006

By Carolyn Kenney, Jianjun Liu, Zhenghe Xu

"The flotation of fine particles has been a challenge to scientists and engineers in mineral processing. A novel concept of reactive oily-bubbles as carrier to float fine particles is described in this paper. The reactive oily bubble flotation is considered attractive for fine particle flotation. In addition to the role of fine particle agglomeration by oily films, the surface properties of air bubbles coated with a thin oil film can be better controlled for the desired selectivity by adding certain types and concentrations of water insoluble collectors into the oil phase. This approach can eliminate the addition of collector to the aqueous phase, avoid undesired synergetic interactions among collectors, activators, depressants and dispersants present in slurry, minimize undesired activation of gangue particles and significantly reduce the amount of collectors needed.The concept of using reactive oily-bubble to achieve selective flotation was demonstrated in contact angle measurement and by microflotation tests with fine sulphide minerals: sphalerite and galena in this paper. The flotation behavior of sulfide minerals can be radically modified with adjustment of solution pH, addition of depressant and activator as in conventional flotation. The successful separation between PbS/SiO2, ZnS/SiO2 and PbS/ZnS was demonstrated using mixed mineral systems. Compared with conventional flotation, the reactive oily-bubble flotation showed a stronger collecting power.INTRODUCTIONThe challenge for fine particle flotation has long been recognized and well documented (Fuerstenau, 1980). Because of drastic changes in surface properties of fine particles and associated physicochemical and hydrodynamic conditions of flotation systems, fine particles responded poorly to the conventional flotation processes. For instance, the small mass and hence low inertia of fine particles would result in a low efficiency of collision and attachment of fine particles with flotation bubbles. Large specific surface area of fine particles, on the other hand, would cause an increased reagent consumption. Dominating colloidal forces of fines may cause slime-coating, accompanying a low selectivity (Somasundaran & Arbiter, 1979; Fuerstenau, 1980; Sivamohan, 1990)."

Jan 1, 2003

By Daven N. Compton

The optical properties of Au nanoparticles have been known for some time, with both Au colloids and Au nanoparticles embedded in glasses having been extensively studied. In this study, Pt nanoparticles have been incorporated into Au colloids, and the linear and non-linear optical properties investigated. UV-VIS spectrophotometry, fluorescence spectroscopy and the z-scan method have been used to examine the effects of such additions.

Oct 1, 2003

By W. Duyvesteyn, B. Sabacky, T. Spitler, D. Verhulst

"The Altair TiO2 pigment process is based on hydrochloric acid leaching of ilmenite at high-acid concentration. Titanium is transferred by solvent extraction into a high-purity aqueous stream with a high titanium concentration. HCl gas is recovered by pressure-swing distillation and is used to regenerate the leaching solution.A new method of TiO2 recovery in dense film with simultaneous dehydration and hydrolysis provides great flexibility in adjusting the characteristics of the TiO2 pigment product.Modifications of this method lead to a flexible, low-cost, high-tonnage process first for the production of nano-sized TiO2, and later for stabilized zirconia-based nanomaterials and nano-sized titanates.IntroductionProcesses for the production of TiO2 from ore by a hydrochloric acid route have been proposed in the past. Most of these processes dissolve impurities and recover TiO2 as a residue or as a hydrolyzed phase in the original solution (Turner, 1948; Berkovich, 1975; Morris, 1982; Judd, 1986; Chao, 1993). Solvent extraction of titanium after leaching in aqueous HCl has also been studied (Tsuchida et al., 1982; Narita et al., 1983) or at least mentioned as an alternative (Ellis, 1963). A number of small-scale studies on leaching and solvent extraction in both sulphuric and hydrochloric systems have appeared recently (John et al., 1999; Lanyon et al., 1999; Sole et al., 1999. As far as we know, only Kerr-McGee (Morris, 1982) and a plant in India and one in Malaysia use a hydrochloric acid route (20 wt% HCl) to commercially produce synthetic rutile.The Altair process originated in a search for an economical method to treat high-magnesium ilmenite deposits, while avoiding the production of a waste metal salt by-product. The first attempts aimed at producing pure TiO2 by leaching out the impurities. Further work centred on completely dissolving the ilmenite in concentrated hydrochloric acid and reprecipitating TiO2. To decrease the impurity level in the product, solvent extraction of titanium into a purified stream was introduced. It also became clear that to take full advantage of the increase in leaching efficiency provided by concentrated acid, it was necessary to regenerate HCl gas."

Jan 1, 2002

By A. W. Stradling, M. J. Fairweather

"The selective flotation of lead, zinc and iron sulphide minerals is key to the economic performance of concentrators treating zinc-lead ores. Diagnostic metallurgy is a tool to quantify mineral behaviour as a function of particle size and liberation. The result of such analysis may be used to select preferred strategies to improve metallurgical performance. The zinc flotation circuits at three Teck Cominco operations: Sullivan, Polaris and Red Dog are compared here. This paper analyses operating data from each of these circuits. The effects of mineralogy and performance by particle size are emphasized. Metallurgical performance at these three concentrators varies substantially. Liberation size is the major factor in determining metallurgical performance. Complex mineralogy requires a fine grind to achieve effective liberation. Fine grinding results in offsetting decreases in concentrate grade through entrainment. The economic treatment of complex fine-grained deposits requires optimisation of regrind size for liberation without unduly compromising concentrate grade through gangue entrainment.INTRODUCTIONUntil recently Teck Cominco Metals Ltd. operated three mines treating zinc-lead ore. In order of increasing mineralogical complexity they are: the Polaris Mine on Little Cornwallis Island in Nunavut in the eastern Arctic, the Sullivan Mine in southeastern British Columbia and the Red Dog Mine in northwestern Alaska. Both the Polaris and Sullivan Mines have recently ceased to operate due to depletion of ore reserves.The ore type and associated gangue, mineralogy and iron solid solution vary significantly between these deposits. Table 1 summarizes some of the important differences"

Jan 1, 2003

By Claude Bazin, Patricia Obiang

"The slurry density (pulp percentage of solids) and the size of the grinding media are often used for the optimization of an operating ball mill. In many cases a change of ball size is made at the slurry density used with the original ball size. But is the optimum slurry density changing with grinding media size? Laboratory grinding tests were conducted to answer that question and results indicate that more than likely there is no need to adjust mill slurry density when adjusting the size of a grinding media in a ball mill. Full scale behaviour could be different because slurry density affects the slurry transportation through the mill and the classifier operation.INTRODUCTIONOre preparation begins with a fragmentation step that liberates valuable minerals or exposes mineral surface to the environment. The size reduction step of a mineral processing operation is energy demanding and the optimisation of this process should always be an objective for the plant metallurgists. Various approaches could be considered for improving the operation of a grinding circuit. For a close grinding circuit an adequate tuning of the classifiers operation may lead to substantial savings (McIvor, 1988). Once a grinding mill is in operation, few variables remain to be considered for process optimization. The ball charge is usually maintained to achieve a target power drawn by the grinding mill motor. The rotation speed is seldom considered (McIvor, 1983; Bazin and Lavoie, 2000) as a change of mill speed usually implies a costly pinion change. If the liner and lifter configuration is critical for SAG and AG mills (Latchireddi et al., 2005), few discussions are available for ball mills (Fuerstenau and Abouzeid, 1985; McIvor, 1983). The mill feed size distribution can also be adjusted but this implies changing the upstream processes, which transposes the optimization process to another level. Modification to grinding media shape is a possible avenue (Cooper et al., 1994), but it remains that slurry density and grinding media size are probably the most frequently considered variables for the optimization of a ball mill."

Jan 1, 2007

By Bob Robinson

In 1990, the European Space Agency (ESA) initiated a technology transfer program to stimulate the wider use of technology developed by the European and Canadian space communities. This paper focuses on the method of technology transfer from the space industry to terrestrial industries, with a focus on the mining resource sector within a program called the Harsh Environments Initiative (HEI) created in 1997. Other resource sector applications include forestry and oil & gas. C-CORE in partnership with the Mining Innovation, Rehabilitation and Applied Research Corporation (MIRARCO) supported by ESA, Canadian Space Agency (CSA) and European R&D organizations conduct projects to promote and demonstrate the transfer of innovative technologies from space to non-space industries. Space programs have and continue to develop new knowledge for extra-terrestrial application, e.g., satellite earth observation, manned missions and unmanned planetary exploration. Many of which hold the potential for application on land and ocean. Currently, space technologies of interest to the mine sector range from robotics and image analysis to sensors, software and advanced materials. The HEI works to identify and apply these types of technologies in the areas of automated/tele-controlled mine operations and mine monitoring.

May 1, 2001

By Marc C. Bétournay

Under the requirement to define the economic advantages of using hydrogen fuelcell vehicles, a broadscoped project was initiated to quantify the costs associated with operating fuelcell vehicles and diesel vehicles. While ventilation savings are clearly an advantage for the former, a complete and detailed analysis of this and other operational and capital cost elements was undertaken in order to provide the mining industry with factual information in considering the changeover from one technology to another. This article outlines the results of the investigations to date defining the costs/benefits associated with: capital purchases, operations (electrical/diesel, mine air heating fuel, ventilation, openings, surface and underground infrastructure, maintenance, labour, etc.), delivery of energy, regulatory obligations, and mine production. Considerations are given to the type and size of a mining operation, the depth, and location in North America. Significant ventilation power, equipment and infrastructure savings are possible with the application of fuelcells for shallower mine applications. At depth, where heat is an issue, there is a reduction in benefit. Specific qualifiers such as minimum air velocities, dust dilution, shaft and heating system velocity limits must be considered on an individual mine basis. While the capital cost for fuelcell mining equipment is currently higher than diesel versions, the high cost of fuelcells is anticipated to be reduced significantly in the near future with the intensification of manufacturing. In either case the operational savings would probably be sufficient to balance the supplementary capitalization costs. A comprehensive cost-benefit report will be issued in the near future.

May 1, 2003

By Sami Hindstrom, Travis Orser, John Starkey

"A great deal of effort has been expended in the last 11 years by the authors and others to identify the correct power required to grind ore in a SAG mill, by developing ore hardness variability functions for the ore body to be mined. This paper describes the next step in the SAG design process, which is choosing a SAG mill to accomplish the required specified design performance.At the test work stage, each ore hardness data point is a unique and discreet bit of information that represents the energy required to grind the ore represented by that sample at optimum SAG conditions. When the design criteria are set, it is now considered necessary that the required power draw in the mill chosen be provided at optimum conditions as well.Based on actual 3-ft diameter pilot plant results, we can now show that the grinding efficiency and throughput drop when the optimum load limit is exceeded. Future work is expected to show a similar reduction in efficiency when exceeding optimum speed. Providing specified power by exceeding these optimum conditions is not adequate and leads to a lower cost but underperforming SAG mill.It is possible in some cases to modify a SAG circuit to force it to perform better, but this often takes time, much effort and costly modifications. In most cases the capital spent on a properly sized SAG mill may prevent production losses that would be many times greater over the extended start up period. This is the direct result of choosing the proper SAG mill at the start.INTRODUCTIONThe sequential steps required for the design of any grinding circuit, have changed dramatically in the last seven years. Previously, it was considered necessary to do six foot diameter SAG mill pilot plant tests on representative samples of minus six inch ore (F80), in order to understand design requirements. It can now be shown that this method is costly and not likely to produce the desired result because it is difficult to recover six-inch material from deep within a new mine. The representative nature of the sample is often compromised when the sample is selected."

Jan 1, 2003

By Charles Kocsis

Over the last few years a multi-faceted feasibility study has been evaluating the introduction of hydrogen fuel cell powered equipment into underground mines to replaced diesel engine powered equipment. The analysis has shown that under normal operating conditions fuel cells could at the very least be an environmental and health benefit by eliminating combustion engines and their unwanted by-products. However, they could also generate significant reductions in the amount of ventilation a mine needs to supply when compared to dieselized operations; this would further decrease a mine?s energy consumption, its associated greenhouse gas impact and also reduce a mine?s significant ventilation associated costs. Fuel cell powered equipment also has the advantage of lower noise and heat production compared to their equivalent diesel powered counterpart. The degree to which fuel cells can generate ventilation benefits in underground mines depends upon a number of operational parameters and mine specific qualifiers such as current ventilation control or management, mine depth, inherent dust conditions and minimum velocity requirements. However another major consideration in the introduction of fuel cells in underground mines will be the safety requirements associated with diluting and removing the potentially explosive hydrogen in the event of a gas leak. In outside surface applications, it is possible to use dispersion, buoyancy, time, and lack of ignition sources as mitigating factors in deriving a low risk of an explosion. However, in underground operations these factors change, dispersion and buoyancy are limited; time may not be available and in non-coal mines ignition sources are not typically controlled. When a hydrogen leak occurs, the availability of sufficient ventilation to dilute and effectively remove the gas will be critical, consequently ventilation may be the prime risk controlling factor. In this paper, the authors evaluate the potential benefits of replacing diesel engines with fuel cells in powered production equipment, discuss the mitigating qualifiers that could limit ventilation savings and evaluate solutions to retain and maintain an additional ventilation capacity in the event of an emergency situation such as a hydrogen leak from the fuel cell stack or its distribution system.

May 1, 2007

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

By J. A. Vallejos

Re-entry protocols are a tactical approach for controlling risk after blasts or large events in seismically active mines. Based on the results of a survey on 18 seismically active mines, mostly in Ontario, current re-entry practices are summarized. Over half of the mines in Ontario use event decay rate analysis as their primary re-entry decision making parameter. We study in detail the applicability of the Omori?s law as an equation for describing the event decay rate for more than 250 mining-induced aftershock sequences in seven different environments. Practical guidelines for the use of the Omori equation as a re-entry criterion are proposed.

May 1, 2009

By Michel Aubertin

Les haldes à stériles contenant des minéraux sulfureux réactifs constituent un milieu favorable à la production du drainage minier acide (DMA). Au sein de ces haldes, les processus physiques, géochimiques et biologiques qui mènent à la production du DMA sont nombreux et complexes. Un des aspects critiques qui affecte la génération de DMA est la distribution et le mouvement de l?eau dans les roches stériles. La modélisation numérique est un outil qui permet de comprendre, pour éventuellement contrôler, les mouvements de l?eau (et donc réduire la production du DMA) dans les haldes pendant la période d?exploitation de la mine et à sa fermeture. Plusieurs simulations numériques des écoulements non saturés dans des haldes ont été effectuées. Pour ces simulations, on a considéré divers types de matériaux, diverses conditions atmosphériques et diverses configurations géométriques. Les simulations réalisées montrent que l?on peut contrôler les mouvements de l?eau en introduisant des couches de matériaux fins intercalées dans les haldes formées de matériaux grossiers. Des couches horizontales ne sont cependant pas appropriées puisqu?elles peuvent, dans certains cas, favoriser la création de zones humides susceptibles d?amplifier les effets du DMA. Une faible inclinaison des couches de matériaux fins vers l?extérieur permet de palier à ce problème. Les couches inclinées favorisent un drainage des haldes vers la périphérie. Ceci a pour effet de garder sèche une grande partie des stériles et de réduire, par le fait même, la production du DMA.

Nov 1, 2002

By Pierre Gaucher

It is possible to selectively mine an orebody by combining the use of the SSW probe with any bulk mining methods as well as narrow vein exploitations. In other words, thanks to the use of the SSW probe, mining operators will be able to reduce and control dilution that is generally associated to every mining operation. Already, quick and substantial savings have been obtained. Falconbridge mines have demonstrated in Sudbury and Raglan operations, that the SSW probe can log nickel, copper and iron sulphides in blast holes. They have raised the grade and reduced dilution by outlining the limits of the ore zones before selectively loading explosives in the blast holes. By lowering the probe in a blast hole, a profile of the ore content is instantly measured and recorded. Logging cored diamond drill holes at the mine have demonstrated that the ore content (nickel, copper, cobalt, PGM) as assayed is proportional to the intensity of the probe?s electromagnetic response (EM), probably because the nickel / pyrrhotite ratio is uniform within a given orebody. The probe also measures the intrinsic conductivity of the sulphides and the percentage of magnetite. All probe readings are displayed and automatically stored at a 10-cm interval within the same reading unit. A 30-meter blast hole takes less than 2 minutes to survey. The whole instrument is light (12 kg) and very simple to use. One operator can carry the SSW system and survey either down or up-holes. The probe works from ?40°C to +50°C. Plotting the measurements of the SSW probe in blast holes (2-3-4-6-10 inches in diameter) before blasting allows to determine the exact shape of the orebody within the waste rock and uneconomic low-grade ore. Thus, only the portion of the holes that contains ore is blasted. The dilution is then much reduced and, by leaving the waste and low-grade ore in place, the mine saves on transportation, blasting, crushing, backfilling, milling, etc. Thousands of dollars are thus saved. Some savings may also be achieved during the development of stopes by guiding development drifts and by replacing diamond drilling by percussion drilling. In some circumstances, when the probe shows that some of the rich ore was not drilled-off, additional blast holes will allow to recover pockets of rich ore that otherwise would have been abandoned in the walls of a stope.

Oct 1, 2001

By N. I. Basuki, E. T. C. Spooner

"Abstract - More than 2800 published fluid inclusion data (primary inclusions) from eighteen typical Mississippi Valley-type (MVT) Zn-Pb deposits and districts were examined to study possible relationships between MVT fluid properties and metal tonnages and ore grades. The mean Th and salinity values for fifteen MVT deposits and districts are 122°C ± 21°C and 20.7 ± 2.6 wt% CaCl2 equivalent, respectively. In general, temperatures increase from pre-ore to ore stages, and salinities decrease from ore to post-ore stages. From post-ore to late calcite stages, salinity and temperature decrease with various trends within specific districts.Several conclusions are drawn: first, results show that Hanor’s (1996) salinity threshold for metalrich basinal fluids applies to typical MVT ore fluids (~16 wt% CaCl2 equivalent). Second, the lower limit of mean Th values of measured MVT deposits and districts (n = 15) of ~75°C is similar to Hanor’s (1996) observed temperature threshold for metal-rich basinal brines of ~60°C (n = 224). Third, there is no statistically significant relationship between fluid inclusion homogenization temperatures (Th) and metal tonnages or ore grades. Fourth, there seems to be a tendency for large tonnage districts/deposits to be associated with mean fluid salinities that are relatively low (~16 to 21 wt% CaCl2 equivalent) within the overall range of salinity means (~16 to 26 wt% CaCl2 equivalent). There is also a correlation between higher ore grades and higher salinities. Finally, comparison with other sedimentary rock-hosted Zn-Pb deposit types shows a trend of salinity decrease and temperature increase from MVT, sandstone-hosted and diapir-related Pb-Zn deposits, to vein-type and Irishtype Pb-Zn deposits, and then to sedex deposits. However, more data are needed for sedex deposits. The salinity threshold for significant Zn-Pb transport decreases to =~10 wt% CaCl2 equivalent at ~250°C. This study suggests that fluid inclusion analysis during exploration may be useful in assessing ore-forming potential (threshold salinity) and possible ore grades (salinity)."

Jan 1, 2002

By F. Flynn, P. Downey, G. Ritson

"Bell Creek Mill which has been in continuous operation since 1987 has undergone numerous operational changes in the past three years. The addition of over 2. 5 million tons of resource at over 13 git from Hoyle Pond Mine precipitated an expansion program for the Bell Creek Mill. This paper discusses the major changes that have occurred over the past two years to increase throughput and efficiencies. In addition, the refractory ore treatment options are also discussed.IntroductionThe Bell Creek Mill is located in the Timmins Goldcamp approximately 8km NE of porcupine Ontario. (Fig I.)The mill was originally constructed by Camamax in 1987 to treat the Bell Creek ore at a design throughput of 350tpd. The flowsheet (Fig 2.) consisted of two stage crushing, grinding, thickening, leach and drum filtration with gold recovery by Merrill Crowe.As ore depleted at Bell Creek, the Mill was purchased by Falconbridge to process ore from their Hoyle Pond property. Kinross purchased the Hoyle Pond property and Bell Creek Mill from Falconbridge in 1992.The successful exploration program at Hoyle Pond has led to several expansions and process design changes, with further changes planned over the next twelve months.In 1994, the mine produced 61,865 oz. of gold at an average cost of 197US$ per oz. In 1995 production was 91, 611 oz. at a cost of 187US$ per oz. Projecting the nine months ended September 30, production for 1996 will be 155,000 oz. at a cost of 165US$ per oz.In early 1997, the CIP expansion will be complete, with a target of 175,000 oz. at a cost of 176US$ per oz for 1997. The mill throughput will have increased from 400tpd to 1500tpd between 1994 and 1997.The changes to Bell Creek since 1994 have been numerous with very little of the existing equipment still in operation. The total cost of the expansion to date is 15 million C$. This paper discusses the major projects carried out during in this time. The focus will be to highlight the justification of each project, the design and engineering, some of the initial start-up problems and solutions.Figure 3 shows a plan of the existing mill facility together with the various additions that were carried out."

Jan 1, 1997

By A. P. E. Dirige

Extensive field study has been performed to assess the capabilities of thin sprayon liners (TSLs) and conventional spray-on support systems for preventing rock and support material damage that often results due to mining-induced rockbursts. Support performance was studied using field scale explosive detonation trials to simulate dynamic failure effects that are known to develop during typical rockburst events. Multiple seismic and high speed photographic monitoring techniques were used to provide detailed information concerning rock motion, surface fracturing, ejected fragment motion, and support liner survivability characteristics. Results of this study have demonstrated that TSL?s and variant layer combinations may be as effective as or better than conventional support materials for mitigating rockburst or like damage in highly stressed mine environments. In this study, field results have been verified using numerical modeling procedures to better understand the support behaviour of TSLs when subjected to highly stressed mine environments and mining-induced rockbursts, a factor that is essential when designing any rock support system. A series of numerical modeling assessments were conducted using FLAC3D modeling of a half circular TSL-lined tunnel, influenced by anisotropic stresses and simulated mining-induced rockbursts, to investigate the support potential of TSLs under prototype conditions. The effect of a mine rockburst was simulated using the three-dimensional dynamic analysis option of FLAC3D. Results of TSL-lined tunnels in highly stressed and mine rockburst conditions indicated exceptional effectiveness of TSLs for suppressing rock deterioration resulting from stresses and rockbursts. The rock support potential of TSLs on tunnel surfaces was compared with that of thin shotcrete linings, and results indicated that thin spray-on lining products currently available may be equally as effective in support as shotcrete materials.

May 1, 2009