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By Garry Johansen, Andrew McCallum, Sandy Gray, Gary Patrick, Jennifer Abols

"The authors challenge the traditional thinking that whole ore cyanidation followed by Carbon in Pulp (CIP) or Carbon in Leach (CIL) should be applied as a standard process in the majority of gold processing flowsheets. The concept of whole ore leaching by using cyanide can result in unnecessary environmental and capital risk. Alternative, currently available technologies can be used to provide the desired level of gold recovery at a much lower installed cost and less impact on the environment. Gravity concentration, alone or in conjunction with flotation, can produce a gold bearing concentrate that can then be fed directly to a batch or continuous intensive cyanidation unit.The InLine Leach Reactor (ILR) is an intensive cyanidation unit developed by Gekko Systems of Australia. Advantages of this technology include significantly reduced capital outlays for greenfield plants or plant upgrades and increased investor returns. Reduction of onsite cyanide and carbon inventories improves plant safety, materials handling and site security. By using direct electrowinning to treat the gold bearing solutions produced by the ILR, the need for a conventional carbon stripping circuit is removed. Potential also exists to increase grind size – saving energy, further reducing capital and operating costs and decreasing the environmental impact of a mill. Detoxification of cyanide before the release of solution to the tailings dam can minimise tailings disposal and containment requirements.When used in the right application the proposed process will have little or no impact on gold recovery and may actually increase recovery for complex ores. Areas applicable to this type of treatment include: plants in environmentally sensitive regions, lower tonnage higher grade deposits, coarse gold deposits, and preg robbing ores - some of which may not currently be economically feasible using standard processes and flowsheets. We have provided three examples of such projects, where an alternative to whole ore leach and CIP or CIL has delivered these benefits."

Jan 1, 2003

By G. D. West, J. D. Purdue

"Marked Ball Wear Testing (MBWT) employs identified grinding balls which are placed in an operating mill for a period of time, retrieved from the mill, and then measured to determine the amount of wear experienced. This is an excellent method to evaluate the relative differences in the wear rates of various grinding ball materials and/ or manufacturing process variables simultaneously and under actual production milling conditions. This paper outlines the methods of ball identification, testing and evaluation procedures and discusses the strengths, weaknesses, and proper application of marked Ball Wear Test results. INTRODUCTIONThe consumption of grinding media in the mineral processing industry represents a substantial cost to the comminution process. Grinding media is most often found to be the second highest consumable cost next to energy. To determine the selection of the most cost effective media, one must determine the specific wear mechanisms present in a given grinding application and decide which media system addresses these mechanisms most economically.A MBWT has definite advantages over other methods used to evaluate the relative consumption rates of grinding media. It allows the comparative testing of a large number of different ball types under identical operating conditions in a very short period of time. In addition, production wear rates can be accurately predicted directly from MBWT results. Typical MBWT test periods are in the order of 350 to 700 operating hours for high wear rate application and upwards of 2000 or 3000 hours for low wear rate applications. These times are much less than the 5000 or more hours required to purge a mill of the incumbent ball charge (to convert it to the test ball) and then provide even a minimal test period in which to conduct a full production test, for just one type of ball. Under production test conditions the evaluation of competitive product or the development of improved grinding balls is necessarily slow."

Jan 1, 1991

By Bern Klein, Malcolm Scoble, Justin Widdifield, Andrew Bamber

"It has been suggested that the pre-concentration of ore underground can bring positive economic and environmental benefits to underground mining, particularly at extreme depth. The integration of mineral processing systems for coarse waste rejection and disposal into the underground mining environment has a substantial impact on the economics and operation of the mine. A case study for the integration of underground pre-concentration at Falconbridge’s Thayer Lindsley and Fraser Copper operations in Sudbury, Ontario is discussed, with particular focus on ore evaluation, system design, integration, and impacts on the operation. Conclusions as to the overall feasibility and potential impacts of underground pre-concentration are presented.INTRODUCTIONObjectivesA scoping investigation has been conducted for Falconbridge Ltd on the Fraser Copper Zone and Thayer Lindsley mines in order to develop a better understanding of the operations, and to do a preliminary evaluation of the potential for pre-concentration of the ore underground. The objective was to develop a conceptual design and evaluation of a pre-concentration system for each site. The results are to be used as the basis for a more detailed future study.Underground Pre-ConcentrationThe pre-concentration of ore to reduce the quantity and improve the quality of material transported to the processing plant is not a new concept and has been in practise at several mines including the American Zinc Company in Mascot Tennessee (Munro et al, 1982), the Sullivan Mine in British Columbia and presently at Xstrata’s Mt Isa lead-zinc mine in Australia, as well as at several PGM mines operating in the South African Bushveld. For each of these mines, a high degree of waste rejection was achieved at a coarse particle size with good metallurgical recoveries. The benefits arising from the surface pre-concentration of ore are well documented in the literature (McCullough et al, 1999; Miller et al, 1978; Munro et al, 1982; Schena et al, 1990; Sivamohan & Forssberg, 1991)."

Jan 1, 2006

By Jim Truskoski, Alain Bouchard, Andy Kerr, Germain Labonté, Jim Barrett

"Improving mill grade recovery relationships is always good business. INCO’s Clarabelle Mill has just undergone a $16 million process upgrade that was designed to improve nickel recovery by 2%. The “Mill Redesign Project” was completed on time, under budget and delivered the promised improvement in recoveries.This paper describes the methodology of the Mill Redesign Team, details of the process improvements and the critical success factors that helped the team deliver on its promises.With a payback of under one year, the Mill Redesign Project will play a key role in ensuring a competitive edge as INCO embarks on its second century of operation.INTRODUCTIONIt is not uncommon in milling operations to encounter a crisis. That is the point when the process does not behave as expected and fails to respond to the increasingly desperate efforts of the metallurgists and operators. Flowsheet and process changes are attempted and abandoned. With luck, one or more of the changes achieves some mitigation of the original problem or more usually, the original cause of the problem disappears. Regardless, the metallurgist can sit back and bask in the apparent success of his or her endeavours.Less common is when a problem is carefully diagnosed, a team put together to explore all avenues and recommend a path of action that is implemented on time, on budget and immediately delivers on the promises which are sustained over time. This is one such story. It is the story of the Mill Redesign Project at INCO’s Clarabelle Mill in Copper Cliff, Ontario.BACKGROUNDThe Clarabelle Mill was built in 1971 on the southern rim of the Sudbury basin. At that time it was one of four mills operated by INCO in the Sudbury area. During the 1970’s, 80’s and 90’s, the need for increased process efficiency, higher levels of sulphur rejection to tails and the concurrent development of high capacity mills and flotation cells, resulted in the consolidation of INCO’s Sudbury area milling operations into a single, expanded facility at Clarabelle."

Jan 1, 2003

By D. B. Felsher

"Lakeshore Gold’s Bell Creek Mill is located in Timmins, Ontario. The mill has been in operation since 2009 and has undergone several small expansions since then. In 2012 - 2013 the first major mill expansion was commissioned raising the mill capacity from 2,000tpd to 3,150tpd. This was done through the installation of a new ROM ore truck dumping facility, jaw crusher, ore storage dome, SAG mill, leach tank, CIL tank, and electrowinning cell. This paper will cover the equipment installed to raise the tonnage as well as the commissioning of the new equipment including problems encountered and how these problems were solved.IntroductionLake Shore Gold acquired the Bell Creek Mill, along with the Bell Creek Mine, from the Porcupine Gold Mines Joint Venture, a joint venture between Goldcorp Canada Ltd. and Kinross Gold Corporation, in December 2007.The Bell Creek mill, which processed ore from the Bell Creek Mine has consistently achieved recoveries of over 95%, is a conventional gold mill circuit, involving crushing and two stage ball mill grinding, gravity and leaching, followed by Carbon-In-Pulp (CIP) process for gold recovery. The facility had been placed on care and maintenance in 2002. At the beginning of 2009, the mill had been refurbished with an initial capacity of 800 tonnes per day and was subsequently refurbished to a capacity of 1,500 tonnes per day near the end of the third quarter of that year. The Mill was then expanded in 2010 to a capacity of 2,000 tonnes per day.In late 2011, the Company commenced a 50% expansion of its milling facility, from a processing capacity of 2,000 tonnes per day to 3,150 tonnes per day. Phase 1 of the expansion, which mainly involved enhancements to the solution circuit (“back end”) and increased the mill’s capacity to 2,500 tonnes per day, was completed at the end of 2012. Commissioning of Phase 2 of the expansion, to a capacity of 3,150 tonnes per day, commenced at the end of July 2013."

Jan 1, 2015

By Mike Ounpuu

"Locked cycle testing has become more common practice in the last 10 years. Reasons vary from a desire to avoid pilot plant testing, demand for more variability testing within deposits, and perhaps most importantly, a request from the financial institutions who invariably finance many of the projects. Surprisingly, many people don’t understand locked cycle testing, analysis and balancing. This paper provides a brief overview of locked cycle testing, and then progresses into a more detailed discussion on the analysis and balancing of locked cycle test data. The paper endeavors to educate CMP personnel on how to determine if the test was good and if a reasonable metallurgical projection was made. The paper uses case studies to illustrate the key points.INTRODUCTIONA locked cycle test is a repetitive batch used to simulate a continuous circuit. The basic procedure has a complete batch test performed in the first cycle which is then followed by similar batch tests which have “intermediate” material from the previous cycle added to the appropriate location in the current cycle. These batch tests, or cycles, are continued in this iterative manner for an arbitrary number of cycles. The final products from each cycle, i.e. final concentrate and final tailings, are filtered and thus removed from further processing. At the end of the test, all the products, final and intermediate, are dried, weighed and subjected to chemical analysis. The test is balanced and a metallurgical projection is made.This description can be found in many of our classical textbooks for the mineral engineer 1, 2, 3,4. Unfortunately, none of these textbooks provide any discussion beyond the basic procedure. Often, they provide comments such as in Taggart’ where he states; “It is questionable whether in any case it approximates mill results any more closely than the standard batch test”. It is truly surprising that our classic textbooks promote that locked cycle tests are more art than science, and suggest that they can be of dubious value. None of the textbooks provide meaningful insight or discussion in how to assess:"

Jan 1, 2001

By C. O’Hara, S. Roy, K. Swedburg, J. Katchen

"New Gold Inc.’s New Afton concentrator commenced operation in June 2012, producing a copper-gold-silver concentrate. During the first two years of operation, various optimization efforts have improved mill throughput, recoveries and concentrate grade. Daily mill throughput has increased 20% above the design rate. This was achieved through more consistent instantaneous feed rate and size distribution from the mine reclaim stockpile, higher pebble crusher reliability, changes to the SAG and ball mill grinding media sizes, larger SAG grate openings, reduced SAG discharge screen-deck aperture size and an increase in the maximum ball mill power draw limit. Metal recoveries at New Afton are primarily limited by liberation in the coarser size fractions; therefore, improving the grind size at a given mill throughput has been the main means of improving recovery. The average rougher flotation feed F80 size was reduced from 175µm in Q1 2013 to 165µm in Q1 2014, while the average operating mill throughput increased by 25%. Other flotation circuit changes focusing on recovery improvement include the addition of froth crowders to the rougher cells and the use of a stronger frother. The average concentrate grade increased from 22% Cu in Q1 2013 to 25% Cu in Q1 2014. This was achieved by redirecting the flash flotation concentrate stream to the regrind circuit and by the addition of lime to the cleaner circuit.This paper describes the above improvements made at the New Afton mill. The paper describes the objectives, timing and end result of each area of improvement.IntroductionThe New Afton Mine is situated approximately 10km west of Kamloops, British Columbia, 340km north-east of Vancouver. The mine has a moderate climate, access to transportation and electrical infrastructure, nearby fresh water supply, and close proximity to industrial and mining service businesses. The population of Kamloops is around 90,000 with most mine employees living in or around the city."

Jan 1, 2015

By K. Mikulash

"Agnico-Eagle Mines Limited is located at Joutel in North-Western Québec. The mine is about four miles north-west from the town on the left bank of the Harricana River. The present mining is 1,800 feet deep with an extensive geological development below this level.HISTORY The development of the ore body started sometimes in 1967 and then the operation ceased in 1969 and again reopened in 1973 due to higher prices of gold. During these periods of active development, a considerable metallurgical research has been done by different laboratories and research and design companies.The original capacities of the Mill differed as much as 30%. So did the concept of the ore treatment. The original design of the mill was for 200 TPD of ore with straight cyanidation. The problems with low gold extraction, high reagent consumption and poor filtration resulted in a decrease of the expected mill output to an initial 550 TPD with gold recoveries in the order of 40 to 60%.Although the studies and research work done on the ore varied, all of them had a common conclusion: a very fine grind is needed to expose gold. The mining method (sub-level blast hole mining) now provides the ore that differs from that on which the metallurgical test work was conducted and the original capacity calculations had been done.These variations in the concept consequently led to the conclusion that the size of gold is a determining factor for the process.Kilborn Engineering provided the original design with the Leslie Eng¬ineering completing the design. The metallurgical test work has been done by the Department of Energy, Mines & Resources, Lakefield Research of Canada, Kilborn Engineering and Sidam Incorporated.The new flotation section with secondary grinding has been designed and put into operation by the company in December 1975 with the results yet to be evaluated."

Jan 1, 1976

By Desmond M. Connor

"The State of the EnvironmentRecent data from a State of the Environment report prepared in January 1991 and released in April (see references) , indicates that the physical and biological environment across Canada is deteriorating in some important ways. For example: • emissions and atmospheric levels of CO2 are rising at a rate more than 50% of the U.S. increase; global temperature continues to rise, especially during the last decade;• while S02 emissions have been falling since 1984, NOz has declined only slightly since that time; N02 in Canada rose 43% during 1970-87, compared with a 7% rise in the U.S.;• municipal discharges of phosphorus and BOD (biochemical oxygen demand) to fresh water have been rising steadily since 1983);• pulp and paper mill discharges to fresh and Pacific coastal waters have been rising, 1970-87;• migratory game bird populations (Prairie Mallard, Northern Pintail) have been declining for the last 20 years;• total per capita energy use continues to climb.The State of the PublicSimilarly, data from recent public opinion surveys shows that public concerns about the environment are either rising, or are remaining at a high level despite the competing claims of the economy and political events for attention.For example, Professor Reginald Bibby's national survey this year shows that Canadians rank our top three issues as theeconomy, the environment and government incompetence.An Angus Reid survey commissioned by the Mining Association of Canada and other mining organizations in 1989 showsthat, despite general support for its economic benefits:• 67% of the general public perceive the mining industry to be a major source of pollution and environmental contamination;• 68% believe that a new mine should not be developed if it will disturb local wildlife;• 52% would prohibit a new mine if it required a road through a national park;• 49% would disallow mine development if it took as much land as an average farm in an agricultural area or if it limited sports fishing in a local lake."

Jan 1, 1991

By J. J. Gillman

"The coal deposits from Cape Perce on Cape Breton Island have been known to supply the coal requirements of the people living in or visiting the area for close to 300 years.A major block of coal has been identified off Cape Perce, which has several seams from which coal can be taken. The initial development by the Cape Breton Development Corporation will be in the Harbour Seam and constitutes the Donkin-Morien Development Project.The CBDC's confidence in this indicated resource is based on its knowledge of the Sydney Coal Field and from boreholes driven into the block during the summers of1977, 78 and 79. In 1980, further geological data were gathered from surface drilling and an intermediate-depth high-resolution seismic survey. The surface facilities for the Donkin-Morien Development Project will be located on the northern half of the cape, with access declines extending 3.9 km in a north-northeasterly direction to the Donkin block of coal reserves. The development has the potential of supplying all the coal needed f or the future electrical power generation requirements of Nova Scotia for many years to come.The planned mine, for which a project definition study has recently been completed, will have a production rate in excess of 3 million tonnes per year.IntroductionAt the Second CIM District 1 Meeting held in Halifax in October 1980, a background paper on Phase II of the Donkin-Morien Development Project was presented. The project involves a 3.2-million-tonne, saleable, underground coal mine. At that time, the chronological sequence of events had unfolded, starting with the initial boreholes through to the feasibility study, and on through site clearing and site preparation. This led to engineering design contracts for the portals and the two tunnels to be driven to intersect with the Harbour Seam close to 4 km from their entry point."

Jan 1, 1982

By Martin Plante, Denis Hamel, Marie-Pier Bédard, Jeff McLaughlin

"The results and methodology for the optimisation of the zinc flotation circuit at the Matagami Mill using the 6-Sigma methodology are shown in detail while only the base line and final results are presented for the copper circuit. As a result of the six sigma project, the zinc concentrate grade was improved by 2% and the zinc recovery by 3%, with reduced dosage of reagents while the copper concentrate was improved by 2.4%. The process variation in the zinc concentrate grade and recovery was also reduced by over 50%. The main factors contributing to the project success were identified to be improvement in the plant measurement systems, an internal benchmarking exercise and statistically designed laboratory and in plant tests to guide the process improvement. The annual value to Noranda resulting from this project has been estimated at approximately $5 million.INTRODUCTIONThe Matagami concentrator started ore processing from the new Bell- Allard deposit in July 1999. Concentrator performance (grade and recovery of zinc and copper concentrates) was highly variable and work was ongoing to fix the sources of variation. In the mean time, Noranda embarked on its 6-Sigma initiative across the company. The 6-Sigma DMAIC (Define, Measure, Analyse, Improve, Control) methodology was used in the two flotation circuits to reduce variation and improve performance. The statistical 6-Sigma methodology moves towards identifying and controlling critical process operating variables (the X' s) to achieve customerspecified performance (the Y's). The goal of the methodology is to describe a process in a numerical equation known as the transfer function, given as Y = f (XI, X2, X3.. .), where Y could be concentrate grade or recovery and examples of X's are reagent dosage, particle size, etc. With the transfer function, optimal operating conditions and control tolerances are established for the key X's to ensure the desired Y performance is obtained.Each phase of the 6-Sigma methodology is explained below with examples from the Matagami project."

Jan 1, 2001

Mining companies around the globe have long struggled with the need to ensure that earth moving equipment is designed to be operated and maintained under all site conditions without causing harm to people. Incident and accident data from a range of sources clearly identify mining equipment related issues as major causal factors. One Australian source, the NSW Department of Primary Industries? International Mining Fatality Database Project Report (2008) noted that ?Approximately 77% of mining accidents resulting in a fatality have? involved equipment?. The largest number of fatalities were associated with trucks; the next with LHDs. Many of the technical advances in equipment design over the past few decades have not been matched by advances in the human factors design of mining equipment, resulting in a ?design vacuum? that must be filled prior to commissioning new equipment. This ?design vacuum? is generally filled by local dealers. They retrofit equipment according to customer requirements to mitigate risks that have not been designed out by the Original Equipment Manufacturer (OEM). Mining companies have been frustrated by lost time and additional costs associated with this need to retrofit ex-factory designs. Some companies have attempted to address the problem by negotiating company or site-wide ?solutions? to equipment designs, but this has not simplified or clarified the issue from the point of view of the equipment manufacturers. A number of mining companies operating in Australia first discussed the concept of a joint customer approach to improve the Human Factors design of earth moving equipment at the factory level in 2004. They agreed that a new approach to engaging with OEMs was needed; one that would define the ?Problem?, in other words, the basis of the risks, rather than dictate the ?Solution?. The Earth Moving Equipment Safety Round Table (EMESRT) was formally established in 2006 by six global mining companies to establish a process of engagement between Original Equipment Manufacturers (OEMs) and mining customers.

May 1, 2009

By Robert B. Parsons

"Starting January 1, 1987 employers, employees, and other taxpayers are confronted with a new set of rules dealing with the taxation of travel and housing benefits relating to isolated posts in Canada, and tax deductions relating thereto. The new rules may not be the simplest in the world, but they are reasonably fair and at least offer a hitherto unknown degree of certainty as to how such allowances are to be treated. IntroductionThe first shoe fell on February 26, 1986 when Finance Minister Michael Wilson' s first budget presented the framework for the future tax treatment of travel and housing benefits (so-called ""northern allowances"") provided to residents in remote regions of Canada. The second shoe fell some ten months later on December 22, 1986 when the Minister of Finance filled in the earlier announced framework with detailed rules. We now have a complete picture of the tax treatment of northern allowances. Here is what the picture looks like with effect from January 1, 1987.According to Revenue Canada, the net housing benefit to employees will be equal to the value of the accommodation provided, less any rent they have paid. In major centres such as Yellowknife and Whitehorse, having a developed rental housing market, the benefit amount will be based on the fair market value of the rental accommodation in that location. In other areas, employers will be required to value the housing benefits and, to assist employers, Revenue Canada has provided the following guidelines: Common shelter- $100 per month including the value of subsidized utilities Apartment! duplex-$270 per month excluding the value of subsidized utilities, or $400 per month including the value of subsidized utilities House-$450 per month excluding the value of subsidized utilities, or $650 per month in- c1uding the value of subsidized utilities.The amounts established as the values of accommodation will be adjusted annually after 1987 based on the housing component of the Consumer Price Index."

Jan 1, 1987

By D. H. Wilton, C. I. Godwin, A. J. Sinclair

"A computer-based drill core logging system used to examine computer-based drill core logging system used to examine copper mineralization hosted in volcaniclastic rocks at the Sustut deposit, in north-central British Columbia, has many advantages over conventional logging. A relatively simple format for recording geological information allows complex, but significant, geologic parameters to be recorded consistently, accurately and quickly. This allows graphical presentations and statistical interpretation of the data that lead to rapid and detailed analyses of the control and distribution of copper mineralization at the Sustut deposit.Ore mineralization is concentrated in lens-shaped bodies and is zoned from native copper cores through chalcocite, bornite and chalcopyrite to a margin containing pyrite. Chalcocite, the principal mineral, occurs mainly in volcaniclastic agglomerate characterized by a specific grain size, a narrow range of matrix and fragment proportions, and a specific sorting.IntroductionWe report here a case history of the practical application of a detailed computer-based drill core logging system applied to the Sustut copper deposit of Falconbridge Nickel Mines Ltd. The Sustut deposit, containing a few tens of millions of tons of ore grading slightly more than one per cent copper (Church, 1975), is in north-central British Columbia about 875 km north of Vancouver. The deposit is contained in volcaniclastic sedimentary rocks of the Upper Triassic Moosevale Formation, which is within the upper part of the Takla Group (Monger and Church, 1977) of basic volcanic and associated sedimentary rocks. Takla rocks occur along the eastern fringe of the Intermontane Belt in the Canadian Cordillera (Sutherland Brown et al., 1971). The deposit consists of concentrations of copper sulphides, native copper and pyrite within pods and lenses parallel to bedding in the host sedimentary rocks. These sheet-like deposits exhibit a mineral zoning from pyrite at the fringes successively through chalcopyrite, bornite and chalcocite to native copper at the core (Wilton, 1978; Wilton and Sinclair, 1979)."

Jan 1, 1980

By D. A. Stokes

"In October of 1980, Fording's 15 000-ton capacity clean coal silo was put into operation. The silo was unique in that it was designed to allow the mass flow of material, coupled with a sophisticated computer-controlled dynamic loadout/weighing system.On May 1, 1982, an explosion, attributed to a methane buildup, devastated the top of the silo. .As a contingent plan to minimize the effect of the production loss the conveyance to the silo top was reconstructed and the silo' was back in operation by mid-June 1982, although the silo top was left open to prevent further methane buildup. As dust emissions occurred during loading ofthe silo, an extensive study was undertaken to establish dust control for the silo, while at the same time providing methane detection and control.The study, conducted for Fording by the Cambrian Engineering Group of Calgary, involved an extensive search and investigation of existing coal silos. The study concluded that there IS no universally accepted method of controlling methane emissions in coal silos.The study also produced a unique conceptual design for handling methane, while at the same time, protecting the environment . Fording's Clean Coal SiloThe silo, located at Fording Coal's Fording River metallurgical coal operation, is utilized for the storage and load out of cleaned thermally-dried metallurgical coal.The silo was constructed in 1979 and 1980 to reduce unit train loadout times and to facilitate the handling of “special shipments"" of coal. The silo went into operation in October 1980, under manual control and utilized computerized, automated loading by February 198I.The silo is 21 m in diameter and 76 m high, with a nominal capacity of IS 000 live tonnes. The coal storage portion of the silo is 21 m in diameter and 48 m high, utilizing 2 steep-angled hopper outlets allowing mass flow of the stored coal during train loading. Each hopper has a control gate and a discharge chute, one chute for bulk loading, the other for trim loading of coal into the rail cars."

Jan 1, 1986

By R. B. Smith, D. C. Brady

"A review of the major renovation of the fifty-year old C-I headframe facilities at NERCQ's Con Mine, Yellowknife, N. W.T. is provided. The project was complete with only two 3-day shutdowns of the facility .Work consisted of four major phases: structural reinforcement of the adjoining mill building prior to dismantling; relocation of virtually all mine and mill utilities which transited the headframe collar en route from the adjacent powerhouse; dismantling of the entire headframe facility and portion of the mill building, including the removal of asbestos siding; and erection of a new headframe, idler rower, collarhouse and mill wall.This paper describes the work and methods used by NERCQ, J.S. Redpath Mining Consultants Limited, Acres International Limited, and others to bring the project to a successful completion. The history of the project is discussed from the planning phase through construction, including; design, procurement, and construction scheduling. The paper discusses the unique elements of the project, the problems encountered, and the methods used to achieve success.IntroductionThe Con Mine is a 800 tpd gold operation located in Yellowknife, Northwest Territories. Originally owned and operated by Cominco Limited , the Con Mine was sold in December 1986 to NERCO Minerals Company, and officially renamed NERCO Con Mine, Ltd. The Con Mine has been in operation for over 50 years. The first gold brick ever produced in the Northwest Territories was poured at the Con Mine in 1938. The original C-1 shaft was sunk in 1937 to a depth of 235 ft. In 1941 the shaft was deepened to a depth of 950 ft, and the original timber head frame was replaced with a steel structure. Following the discovery of the Campbell Shear in 1946, the C-l shaft was again deepened to its present depth of 2443 ft. Over the years, mine production gradually moved deeper and further from the C-l shaft."

Jan 1, 1991

By Rory S. McLaren

"Cost per ton as it relates to mobile underground vehicle maintenance and repair is invariably astronomical. The reason: the mechanic replaced a ten micron filter with a 250 micron filter because it appeared to be physically identical; the purchasing agent bought ""Brand X's"" filter because it was 25% cheaper- the result, a filter life 30% less; the hydraulic pump was sent for repair at a cost of $2000.00 and failed shortly after installation; pumps are failing because there is no testing equipment available in the workshop.This paper will identify a number of the common pitfalls resulting in high costs, and offer solutions to bring these costs to within acceptable limits.The following key issues will be discussed:A. Genuine parts replacement versus ""Brand X""B. Evaluating outside repair shopsC. The influence of environmental conditions on optional equipmentD. Personnel trainingE. Servicing and testing equipmentIntroductionThere is an old English aphorism, ""Look after the pennies, the pounds will look after themselves"".This paper is aimed at the pennies.Genuine Parts Replacement versus ""Brand X""High interest rates and floundering mineral prices are forcing many equipment owners and operators to purchase ""Brand X"" or ""Pirate"" parts in order to survive. ""Pirate"" part manufacturers invariably offer lucrative discounts which are most appealing to the average purchasing agent whose responsibility it is to save money. Unfortunately, in most cases, there is a tradeoff involved; when the price drops so does the quality, which can adversely affect owning and operating costs. Unless totally obvious, ""original"" or ""pirate"" part metamorphosis requires scrupulous analysis by competent individuals.Hydrostatic transmissions have been used successfully for many year s in various industries. They found their way into mobile underground loaders in the early seventies and have since become very unpopular in mining circles due to their apparent ""inherent unreliability"" and near ""prohibitive"" repair costs. The big benefit of the hydrostatic transmission is compact component packaging which enables the machine to work in very small headings."

Jan 1, 1983

By J. Jara

Oxygen, ozone and air were evaluated in the oxidation of ferrous iron in the presence of hydrochloric acid ranging from 0.1 to 4 M HC1. Stoichiometric coefficients were determined through mass balances of gases and ferrous titration. The molar ratio ferrous to ozone increases from two at 0.1 M HC1 concentration to six at 4 M HC1. For oxygen the same ratio is constant and equal to 2 and it is in agreement with other works. The stoichiometric coefficient of hydrogen ion was not defined due to the effect of hydrolysis of ferrous and in particular ferric ion on measuring acid concentration by titration with sodium hydroxide. Agitation, gas flow and pressure contribute in the utilization of gas. A technique to compare the gas efficiency at several conditions is discussed and it is based on the relationship between oxygen and ferric during time. There is a strong compromise between gas efficiency and ferrous conversion. Oxygen efficiency decreased with time due to ferrous depletion. The gas efficiency ranged from 3 % (low level of agitation) producing 25 g Fe(III)/h to 12 % at 1000 rpm with 60 g Fe(III)/h converted. At 405 kPa absolute the oxygen efficiency was in the range of 80 % producing 300 g Fe(III)/h. The replacement of chlorine with oxygen to enrich molybdenite concentrates by reducing copper to less than 0.2 % was evaluated. Oxygen can be employed in batch or continuous operation and it can treat molybdenite concentrates with high variation of copper content. The dissolution rate is mainly controlled by oxygen addition in which utilization of oxygen plays an important role. Improvement on safety and environmental conditions is feasible by using oxygen instead of chlorine gas. The economics indicate good opportunities to implement oxygen without significant modifications to the actual circuit and enough benefit when oxygen is well used.

Jan 1, 2007

By D. E. Wakefeld

"My first reaction to Joanna LaForte’s discussion paper is that she has quite properly assumed that I would like the process of equitable participation between aboriginal landowners and mining companies to be a fast process. On the other hand, I have been a business lawyer long enough to know that the past experiences of the parties will have a significant effect on how they will proceed at any one time.In my article, I was trying to convey the importance of responding to the wishes, aspirations and fears of aboriginal communities in Arctic regions without land settlements. I agree that if it is necessary to wait for governments, the process will be quite slow. My article did point out that often aboriginal communities and mining companies who have become comfort-able that a project is in their mutual interest regard the government(s) involved as their common enemy. I agree that the education and training of aboriginals should “expedite the negotiations and end the power struggles equitably”. Also, I did not mean to suggest that the elders distrust of changes in traditional land use was solely due to concern for alienation and separation from their offspring; however, I suspect it is one reason for their caution. I also agree with this author that establishing opportunities that enable the young aboriginal community members to remain in the north and avoid the problems that await them in southern urban culture should be a common concern of the mineral developer and the Arctic aboriginal community.When this author suggests I am “overly concerned about how much land the aboriginals will control, how educated they may become, how ignorant they are of mining procedures, how objections from aboriginals will escalate expenses, etc.”, she attributes to me as “concerns” matters that were intended to be general information for my readers. My references to the increased land holdings and the improvement in education of aboriginal peoples was intended to deter mining companies from any attempt to be inequitable in their dealings with aboriginal communities."

Jan 1, 1996

By Richard C. Williams, Herman Pieterse, Ken Thomas, K. S. Fraser, Ron H. Colquhoun

"The Banick Goldstrike Mines Inc. (Goldstrike) operation is located in Eureka County in north-central Nevada on the Carlin geological trend. Gold in the sulphide mill feed, occurs primarily as submicroscopic inclusions or possibly solid solution in a pyrite and marcasite matrix. Sulphide mill feed is pretreated in a pressure oxidation plant allowing gold recoveries of 90 per cent and greater to be achieved. Anticipated 1994 gold production for the sulphide gold facility is approximately 1,500,000 ounces.In January 1989 American Banick Resources Corporation (Banick) announced the US$365 million Betze Development Plan for the Goldstrike property. The Betze plan was based on a three (3) staged expansion of the original 4,500 STPD oxide gold plant. This expansion would involve addition of ore and gold processing capacity and the total conversion of all ore handling facilities to treating Betze sulphide ores.Phase 1 of the Betze Development Plan involved the construction of a pressure oxidation circuit, including a single autoclave and supporting ancillary services. The pressure oxidation circuit was commissioned in February 1990. Currently a throughput of 2,000 STPD is being achieved.Phase II of the Betze Development Plan, commissioned throughout April and May 1992, included the expansion of grinding, leaching, and gold recovery facilities enabling the treatment of combined oxide, and sulphide ore. Pressure oxidation capacity was increased by the addition of two (2) autoclave circuits each treating 2,600 STPD of sulphide ore and other required ancillary services. By June 1992, the Phase II gold facility had achieved a combined throughput of 15,000 STPD of oxide and sulphide ore.Phase Ill, the final stage of the Betze Development Plan, was successfully commissioned in February and March 1993. This final stage of development increased the design throughput of the pressure oxidation plant to 15,000 STPD by the addition of three (3) autoclave circuits with combined capacity of 7,800 STPD."

Jan 1, 1994