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By Orlando C. Gaspar

The presence of ?ne-grained mineral intergrowths, the complex distribution of deleterious minor elements, and the variable base-metal contents in some assemblages commonly pose technical problems for the processing of volcanogenic massive sul?de ores. The bene?ciation of most massive sul?de ores from the Iberian Pyrite Belt requires extremely ?ne grinding to liberation sizes of K80= 15 µm to 25 µm. For several deposits, mapping of the ore types and metal zoning pat-terns has shown that the different ore types are distributed as ore lenses within the massive sul?des. Mineralogical, textural, and metallogenetic studies have provided an understanding of the distribution of both the vertical and the horizontal metal zonation in the orebodies. Because the ef?ciency of ore processing is affected by ore textures, and because the same mineral commonly has different textures and varies in chemical composition, a relationship between textures of the feed and the middlings must be established for all ore types de?ned at the deposit scale. Quantitative mineralogical methods have been developed to provide, on a timely basis, the information needed to optimize the concentration processes for the massive sul?de ores at the Aljustrel and Neves-Corvo deposits, Iberian Pyrite Belt, Portugal.

Jan 1, 1997

By Peter Wells, Adrian Sabau, Turgut Yalcin

"The nickel ores of Sudbury contain substantial amounts of the gangue mineral pyrrhotite (Fe7S&) which causes certain diflculties in the milling of these ores. This paper proposes a process called “magnetoflotation” as an effective means of dealing with pyrrhotite. The method reduces the pyrrhotite content of concentrates by applying a magneticjield on the path product as it leaves the flotation cell.Successful performance of magnetoflotation requires the optimization of several variables. In this study, three of the variables (air flowrate, collector concentration, pother concentration) were investigated by Box- Wilson experimental design, and others (magnet speed, grind size, pulp density, flotation time) were examined individually. The process was tested on a Falconbridge nickel ore, assaying 1.92% Ni, and under the best operating conditions, a concentrate was obtained with 13.56% Ni grade and 74.32% recovery in a single step. Further improvements to grade and recovery were possible by cleaning and scavenging. The same ore was also tested with conventional methods for comparison, and results indicated that magnetoflotation performed better.INTRODUCTION“Magnetoflotation” refers to the process of carrying out magnetic separation and flotation simultaneously within the same device, and is intended to prevent magnetic minerals from reporting to the froth products while non-magnetic minerals are being floated. Typical cases where this method might be beneficial include the processing of magnetite and nickel ores. Its use on magnetite ores has been reported elsewhere [Yalcin, 1992; Yalcin, 19931. This paper describes its application to nickel ores, more specifically to the nickel ore of Falconbridge Limited in Sudbury."

Jan 1, 1999

By Harry Siewert

"The commissioning of the 4th concentrator (C4) at Freeport Indonesia in early 1998 increased total mill throughput from 130k to 230k tonnes per day. The C4 flowsheet consists of a single 38 ft diameter SAG mill, four 24 ft diameter ball mills followed by four banks of 4,500 ft3 WEMCO flotation cells.The design of the C4 rougher flotation circuit was based on the expectation that the copper and gold recoveries achieved with the 4500 ft3 flotation cells would be equivalent to or better than that achieved with the 3,000 ft3 flotation cells in the 3rd concentrator (C3) rougher circuit. This assumption was due to the fact that the C4 flotation feed size would be less than 15% +212 microns compared to the C3 rougher flotation feed size of less than 23% +212 microns. By mid 1998 it had become apparent that the C4 rougher flotation copper and gold recoveries were below those predicted.In order to determine the cause of the lower recoveries, stage by stage rougher flotation surveys were performed in all four concentrators. The results of the surveys showed that the C4 rougher flotation circuit had a lower recovery of +200 micron material than that achieved in the C 1, C2 and C3 concentrators. Through mineralogical analysis, it was determined that a majority of the copper contained in the +200 micron fractions was liberated and recoverable. This paper describes the steps taken to increase coarse particle recovery, overall plant recovery and the results achieved."

Jan 1, 2001

By Franciszek Letowski

"Recovery of precious metals associated with a sulphidic matrix has an important economic effect on complex sulphide ore treatment. Because a significant portion of the silver and gold associated with pyrite are rejected to tailings during the selective flotation of pyritic Zn-Pb-Cu ore, other ore processing routes have been investigated.Direct leaching of the ore with ferric sulphate solution followed by separation of different components of the solid suspension by flotation was applied. Such a process allows for: production of zinc sulphate solution; transformation of galena PbS to insoluble lead sulphate PbS04 and its accumulation in a non-floatable fraction together with precipitated geothite FeOOH and other insoluble gangue minerals; separation of sulphur and non-decomposed sulphides in the floatable fraction.Laboratory research carried out on the ore samples from an Alaskan pyritic deposit from the Delta region indicates that gold and silver are mostly concentrated in the floatable fraction.Total recovery of zinc from the ore attains 96.5% to 98.5% after ferric sulphate leaching. About 85% of copper was extracted from the ore during the ferric sulphate and subsequent ferric chloride leaching, and about 6% of the copper remained in the non-floatable fraction and mixed fraction during the leaching-flotation step. Extraction of lead from the nonfloatable fraction during leaching with calcium chloride solution attained 98% to 99%. Extraction of silver from the floatable fraction during leaching with ferric chloride solution reached 98%. Extraction of gold from the floatable fraction during cyanide or thiourea leaching reached 71% to 89%. The results of the laboratory research are presented."

Jan 1, 1987

By Eric Konigsmann

"Omai Gold Mines is a large scale open pit gold mining operation located in central Guyana, South America with a 12,000 tonne per day milling facility. Mill operations were initiated in December 1992 with commercial production beginning in January 1993. The entire operation employs approximately 120 expatriates and some 800 Guyanese nationals and is expected to produce in excess of 250,000 ounces of gold per year. The total mineable reserves contain 2.5 million ounces of gold with a projected mine life of twelve years.The milling circuits feature an SABC grinding circuit, gravity concentration and a conventional CIP gold recovery circuit. This paper will review the mill start up and ""run-in"" phase during the first year of operation. Emphasis will be placed on descriptions of start up problems and circuit modifications, mill performance trends in 1993, and personnel training and development. An outlook for 1994 is also given.IntroductionOmai Gold Mines is a large scale, open pit gold mining operation located in the Potaro district of Guyana, South America, some 160 kilometers south of the capital city Georgetown. Although there are numerous small scale gold mining ventures in Guyana, particularly with the dredging operations on the rivers, Omai is the first hard rock, open pit mine to be developed in the country. At present, this mining operation is one of the largest gold mines in South America.The Omai gold deposit has been exploited for over one hundred years. In the late 1800's, a German syndicate used hydraulicing with subsequent gravimetric separators such as sluice boxes and trommel screens to recover gold from saprolite, a soft, oxidized, claylike mineral which covers much of the claim area. Through the early 1900's and up until the 1950's, numerous mining companies including Anaconda Mining Co. Ltd. explored and drilled the Omai deposit but the technology and low gold price of the day made the reserve uneconomic to develop."

Jan 1, 1994

By A. Bamber

Canadian leadership in a global mining context will depend not only on entrepreneurship and best practice but also on technological innovation. This paper argues for more ambition and commitment to R&D for technological innovation in Canadian mining. It focuses on underground hardrock mining, where past R&D initiatives have simply not stood the test of time, and it is important to examine why. Concerns arise because industry’s commitment to R&D has not been sustained over time. Innovation through exploration, processing, and environmental technologies has been more significant. Meanwhile, R&D capacity and synergies in government institutions and universities have not been exploited. There is a need for more commitment, coordination, and collaboration by industry, governments, and universities. This paper examines some corporate models that have achieved sustainable innovation, along with some that have not. A strategy for innovation is discussed that relates to culture, commitment, coordination, and collaboration.

Jan 1, 2005

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

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

Jan 1, 2008

By J. A. Stewart

"A common problem in the mining industry involves the overloading of established conveyor systems. This overloading can have a number of causes, such as new production technology or a changing market condition which increases demand. An overloaded system usually spells disaster and leads to clean-up problems, equipment failures and lost production. The Cape Breton Development Corporation's Lingan Mine had a serious overloading situation on its underground coal clearance system, and equipment was being damaged and production was being lost. It was rectified, in part, by increasing the coal clearance capacity through the use of booster conveyors. This paper discusses the problems which led to this situation, the booster concept and the installation of the first unit at the Lingan Mine.IntroductionThe Lingan Mine is a submarine coal mining operation with an annual production of 1.5 million tonnes from three advancing longwalls. Coal is conveyed 3400m up a 13% slope to the surface. Designed and installed in the early 1970s, the system's 820-tonne/hour design initially proved adequate; however, by the late 1970sthe system was handling surge loads of 1640 tonnes per hour due to the introduction of new face equipment.The coal clearance system consisted of three 485-Kw, two 210-Kw and one 105-Kw conveyors. Each conveyor was designed for 820 tonnes per hour using 1050-mm, Type 10 belt with a 35-degree troughing profile.There were many effects of the overloading on the system:• Loading beyond the physical capacity of the belt resulted in spillage along the conveyor, which increased operational and maintenance costs.• Overloading of the drive components resulted in premature equipment failures.• Spillage at the drive pulley often resulted in damage to the belt and pulley lagging and, under extreme conditions, the conveyor would require partial unloading before it would operate. This was a particular problem under wet conditions because the 485-Kw conveyor drives were fitted with single-drive drums."

Jan 1, 1982

By L. R. Plitt, A. J. Neale, B. C. Flintoff

"INTRODUCTIONThis is not a technical paper in the normal sense. It is an informal document which is intended to share the results of some recent cyclone research with our colleagues both in operations and in research.PILOT PLANT FACILITIESFigure 1 is a schematic of the closed circuit hydrocyclone pilot plant testing facility at the University of Alberta. In this figure the system is configured for some experimental work on wide angle cyclones which was recently completed. The rig is designed to study cyclones in the approximate range of 10 cm to 20 cm diameter. Notable features include the variable speed feed pump, the magnetic flowmeter and pressure sensors on the cyclone feed line and the travelling launder which permits simultaneous sampling of both product streams.ORGANIZATIONFor organizational convenience the subject matter has been divided into three parts:(1) Peculiar Partition Curves(2) Wide Angle Cyclones(3) Rope DischargePECULIAR PARTITION CURVESMuch of the initial empirical mathematical modelling work on cyclone classifiers dealt with homogeneous ores. A homogeneous ore is one in which all particles have substantially the same apparent density, or where one phase is very clearly predominant. In contrast, a heterogeneous ore is one in which there are at least two different density classes, each making a significant contribution to the ore composition. (Particle shape similarity is always assumed.)One of the problems which can arise when analysing the overall solids partition (selectivity, Tromp, etc ... ) curve for a heterogeneous ore is the peculiar shape which may be observed, when compared to partition curves for homogeneous ores. This is not an unusual occurrence and it has been documented by Lynch and Rao, Finch and Laplante and Larsen and Spring. The purpose of this section is to identify those operating conditions which may give rise to peculiar partition curves."

Jan 1, 1986

By G. B. Bell

"INTRODUCTION On February 1, 1962, after five years of exploration and research, Jones & Laughlin Steel Corporation announced that they would develop the Adams Mine. Plant construction began immediately and in December of 1964 the first shipment of pellets arrived at J & L's blast furnaces in Pittsburgh. J & L continued to operate the property through its wholly owned subsidiary, Jones & Laughlin Mining Company, Ltd., until August of 1971 when it was sold to Dominion Foundries and Steel, Limited of Hamilton. Dofasco engaged Cliffs of Canada, a subsidiary of Cleveland Cliffs Iron Company, to manage the operation of the property.SUMMARY:The Adams Mine is a one million tons per year iron ore pellet operation located in the Boston Township Iron Range, approximately seven miles southeast of Kirkland Lake, Ontario. The property includes 5,300 acres in which eight areas suitable for open pit mining have been outlined. It was estimated that the reserves in these out lined areas are sufficient to produce 30 million tons of high grade iron ore pellets.The ore 'consists of fine grained magnetite with alternating layers of chert and is mined by conventional open pit methods. This extremely hard and abrasive ore averages about 20 percent magnetic iron, and has a ratio of concentration of 3. 9 /1. Presently, mining operations are confined to two pits with a third one being developed for operations to begin in 1973.The run-of-mine ore is crushed and ground and then concentrated by magnetic separation and flotation to a concentrate grade of 68 percent magnetic iron. The concentrate is agglomerated by a grate-kiln pelletizing process and shipped by rail to Dofasco's blast furnaces in Hamilton and to Jones & Laughlin's furnaces in Pittsburgh."

Jan 1, 1973

By Jeffiey B. Austin, Maurice Tagami

"The ores of the Iberian Pynte Belt are typically massive sulphide ores with very fine grained mineral textures. EuroZinc Mining Corporation is currently evaluating one of the massive sulphide deposits located near the town of Aljustrel, in South Central Portugal. The deposit contains significant grades of lead, zinc and silver. The ores are comprised solely of sulphide mineralization consisting of pyrite, sphalerite, galena and some secondary copper minerals. Liberation requirements for these ores are typically 22 to 25 microns. A selective flotation process has been utilized to produce separate lead and zinc concentrates for sale to off-site contract smelters.This paper outlines the process test results and describes the process parameters required to optimize the process parameters. Ths test program has been conducted in support of the re-opening of the facilities in Aljustrel. The 3500 tpd mining and milling facility was commissioned in 1990 and operated for 18 months before closing. The previous facility was designed to produce three products using a selective copper-lead-zinc flotation process. EuroZinc Mining Corporation is proposing to re-open the operation, selectively mining from a lead-zinc phase of the massive sulphide deposit and produce only lead and zinc concentrates.Results are presented and discussed.PROJECT HISTORY AND CURRENT PROJECT OBJECTIVESThe EuroZinc Aljustrel Project is a potential major producer of lead, zinc and silver located in the Iberian Pyrite belt of south-central Portugal. The town of Aljustrel is historically associated with mining dating to the time of the Romans. The area is rich in pyrite deposits, and sections of the pyrite deposits contain significant zones of copper, lead, zinc and silver. Mining in Aljustrel during the last century (1900’s) included a major pyrite mining operation, which produced feed for sulphuric acid production. Pynte mining operated from the 1930’s to the early 1970’s and a short-lived base metal operation from 1989 to 199 1."

Jan 1, 2000

By Charle Pelley, Kosta Fytas, Peter Calder

"A computer model has been developed which is capable of evaluating alternative production schedules (long- and short range) of an open pit mine once the ultimate pit limits have been defined. The model user can choose the best alternative according to cash flow considerations and the particular mining conditions. The model involves a combination of simulation for long range production scheduling and linear programming for short range scheduling. IntroductionMining of mineral deposits in such a fashion so that, at depletion the maximum possible profit is realized, has been a very old problem of the mineral industry. In the past, when high-grade ore reserves were readily available and profits were very high, no question of optimum profitability was raised.However, recently the complexity of both open pit and underground mining has been increased tremendously. The additional problems of continuously rising production costs on one hand and of depressed metal prices worldwide on the other, have been demanding optimum profitabiliy from every mining venture .Particularly in the area of open pit mining, the sequence of ore and waste extraction has now become very important since the size of the pits and of the mining equipment has increased enormously. The process of open pit mining involves a continuous change of the open pit surface and there are many feasible plans of extraction which lead to the same final pit shape (ultimate pit limits). Each of these feasible plans generates a production schedule over the entire mine life, which mayor may not be optimal."

Jan 1, 1987

By C. Stapel, AA Groningen, R. J. De Meijer

"In recent years the Environmental Research Group of the KVI has been developing a number of radiometric techniques that may be employed in mineral sand exploration. These techniques involve: radiometric fingerprinting for assessing sand provenances and mineralogical composition; thermoluminescence for dating and provenance determination of sediments; a towed seabed detector to map seafloors for radiogenic heavy minerals and assessing heavy mineral volumes; and simplified transport mechanism models to explain and predict heavy mineral concentrations.All techniques exploit the property that heavy minerals as a group contain concentrations of naturally occurring uranium and thorium that are a hundred to a thousand times higher than in light minerals. In this paper the physics of a number of methods will be described together with a number of results. Based on these results potential applications to mineral exploration and mining are discussed."

Jan 1, 1994

By M. Leroux, S. R. Rao, J. A. Finch

"Poor Pb-Zn selectivity during Pb flotation in laboratory tests on a Mississippi Valley-type Pb-Zn ore was traced to apH effect. In the pH range, 9.5 to 11.0 sphalerite floated readily without copper activation. Compared to sphalerite flotation with copper activation this procedure gave equivalent selectivity but reduced flotation rate.Possible mechanisms of this sphalerite flotation without copper activation are discussed. Activation by Fe(OH)2 is suggested, the Fe- + ions being derived from superficial oxidation of pyrite in the ore. IntroductionGalena and sphalerite commonly occur together. A typical procedure, using xanthate collectors, is to float the galena first, at pH - 8, followed by flotation of sphalerite after activation with copper (II) sulphate at pH - 10 (to depress pyrite).Flotation tests following this procedure in an undergraduate laboratory course with a Pb-Zn ore from Polaris failed to achieve consistent selective flotation of galena. Occasionally a considerable amount of sphalerite, even though unactivated with CUSO4, floated. This was traced to accidentally high pH. A systematic study was carried out which revealed that sphalerite floated readily in the pH range 9.5 to 11.0 without activation by CuSO4 This paper describes the results of the study and discusses a possible mechanism of this flotation of sphalerite."

Jan 1, 1987

By J. C. Johnson

Underhand mining with cemented mill tailings for backfill has proven to be a safe and effective alternative to conventional overhand mining for rockburst prone mines. The ability of the cemented sandfill to survive a rockburst shock wave without failing has been observed at mines in the Coeur d?Alene Mining District many times in the last twenty years. To better understand the dynamic loading of this commonly used support material the National Institute for Occupational Safety and Health (NIOSH) performed laboratory experiments using a split Hopkinson Pressure Bar (SHPB) to quantify the maximum stress and energy absorption of cemented mill tailings from the Galena mine. Cylindrically shaped cemented backfill specimens aged 28 days were placed in the split Hopkinson pressure bar testing apparatus and a dynamic load was applied. The measured peak dynamic compressive strength of the cemented sandfill was 10.6 MPa (1540 psi) and is approximately twice the static unconfined compressive strength (UCS) of 5.5 MPa (800 psi). These test data showed about 95% of the initial energy was reflected away from the backfill and only 5% of the energy was absorbed. This may help to explain why backfilled stopes could withstand seismic events up to 3.5 Richter magnitude (Ml). When a compressive rockburst wave impacts a stope a large amount of the stress and energy is reflected away. The reflected wave is a tensile wave which creates tensile cracks and spalling in the entire wall of the stope. The backfill supports the damaged wall rock and holds it in place. Some of the stress and energy of the rockburst is transmitted into the backfill and because the backfill is being dynamically loaded it has a strength that is twice its static value and absorbs a substantial amount of rockburst energy. Although mining in deep underground mines is an extremely hazardous occupation, the use of underhand mining with cemented backfill for static and dynamic ground control increases the safety of the workers.

May 1, 2007

By A. Esterhuizen, S. Essack, H. Joubert, N. Fowler

Tenova Pyromet has developed innovative technologies for copper smelting and electric slag cleaning furnaces. Based on client feedback and requirements, innovative designs have been developed to address key design areas and improve furnace reliability, operability and hygiene. These developments include the NovaLance™ lance system to prevent lance bending on top submerged lance furnaces, the SlagFlo™ slag tapping flow control device to better control bath levels and tapping rate for granulation, and an electrode seal for improved sealing directly onto the electrode casing to allow water-cooled electrode components to be located above the furnace roof. The potential application and benefits of these innovative designs are discussed. INTRODUCTION Tenova Pyromet has a long history in the engineering and supply of innovative electric furnace and related equipment for the base metals and platinum industries. Recently Tenova Pyromet has used its furnace equipment design experience to engineer innovative solutions for top submerged lance (TSL) furnaces, and potentially other furnace technologies. These equipment design solutions are developed to address specific operating challenges and operator requirements. Three of these innovative equipment developments are discussed in this paper. First the NovaLance™ lance technology to prevent lance bending during operation in TSL furnaces, such as used for primary copper smelting, is described. Lance bending has been identified as a significant problem for TSL furnace operators as it causes down time and affects furnace availability. The SlagFlo™ device to control the slag flow rate during tapping has been developed based on specific requirements in the mineral wool industry. The SlagFlo™ device can be used to control slag bath levels in primary smelting, converting and slag cleaning furnaces for base metals including copper. In addition, it can be remotely activated for the safe closing of slag tapholes located upstream from a water granulation system. Finally, Tenova Pyromet developed an electrode seal that seals directly onto an electrode surface. This is a special requirement related to base metals and platinum electric smelting and slag cleaning furnaces where furnace operators prefer to keep the water-cooled electrode equipment above the furnace roof. The seal can be used to seal directly onto the potentially hot and uneven surface of Söderberg, pre-baked and graphite electrodes.

Jan 1, 2019

By S. Aram

Product Lifecycle Management (PLM) solutions have been serving as the basis for collaborative product definition, manufacturing, and service management in many industries. They capture and provide access to product and process information and preserve integrity of information throughout the lifecycle of a product. Efficient growth in the role of Building Information Modeling (BIM) can benefit vastly from unifying solutions to acquire, manage and make use of information and processes from various project and enterprise level systems, selectively adapting functionality from PLM systems. However, there are important differences between PLM?s target industries and the Architecture, Engineering, and Construction (AEC) industry characteristics that require modification and tailoring of some aspects of current PLM technology. In this study we examine the fundamental PLM functionalities that create synergy with the BIM-enabled AEC industry. We propose a conceptual model for the information flow and integration between BIM and PLM systems. Finally, we explore the differences between the AEC industry and traditional scope of service for PLM solutions.

Aug 1, 2013

By Raymond Philippe

Background ?Hatch Water has worked with most major mining companies in Chile on water supply and optimization projects ?Many(copper) mining projects are located in Northern Chile, which hosts the driest desert on earth Water supply options in Northern Chile ?River ?Availability/Rights ?Well(Brackish) Water ?Availability/Rights ?Municipal Wastewater ?Water rights ?Volume ?(Desalinated) Seawater ?Infinite source ?Cost depends on distance and altitude ?Recycling/Reutilization ?Water losses All these options present water quality issues

May 1, 2009

By B. M. Hamilton

"IntroductionTHE PROCEDURE and technique for drilling holes has probably undergone as great a change during the past ten years as any one of the other operations conducted in hard-rock mining. During this period, we have seen a change from the heavy leyners and drifters to the light jackleg type of drill machine; from steels that were forged and hardened on the bit end to the inserted •carbide-tipped rod, from the necessity of providing special bit-end fabrication and beat treatment of shank steels to the simple grinding or turning of a plain section on a shank with no heat treatment required for bit attachment.When the steel maker was first called upon to manufacture a hollow drill rod, it was required that the steel must not only have sufficient strength and fatigue resistance to cope with the stresses of service, but also must be capable of being heat treated on the bit end to a high hardness for resistance to wear and abrasion . Little was known at that time about alloy steels and the beneficial effects to be gained from additions of chromium, nickel , and molybdenum. So it was that the plain 0.80 per cent carbon analysis was selected, since it had already proven to be a suitable grade for chisels, hammers, and other such shock tools. This grade has long endured the test of time and is still used in considerable tonnage today, particularly by the construction industry.As the metallurgist learned more about the influence of alloying elements, a new hollow drill analysis was placed on the market in the early 'thirties'."

Jan 1, 1954

By K. Swedburg, M. Samuels

"The New Afton concentrator was commissioned in June 2012, reaching the design milling rate in September 2012. By late 2013, through a campaign of optimization and de-bottlenecking, it was recognized that the installed plant was capable of greatly exceeding the design throughput of 11,700 tpd. Because the value of the metal contained (by increasing the plant feed rate) exceeded the offsetting losses of copper and gold recovery (as the operation moved to a coarser grind), the concentrator feed rate was strategically increased with support of New Gold, while the operations team was given a mandate to evaluate, and then implement, flow sheet options that could re-establish the original design criterion grind target and achieve the corresponding improved recovery. Broadly, the two options implemented were the addition of a tertiary grinding stage consisting of Metso Vertimill® prior to rougher flotation (in order to address the grinding required at increasing throughput by increasing the power), and the addition of three Woodgrove Staged Flotation Reactors (SFR) added to the cleaner circuit after regrind of the rougher concentrate (to address the increased mineral loading to the cleaner circuit). Through test work, it was anticipated that the SFRs would be capable of achieving final concentrate grade in a single pass without the integration of recirculating loads, a strategy sometimes referred to pre-cleaning. This paper describes the expectations generated for the Woodgrove SFR unit operation through pilot plant testing on site, and compares those results with the performance of the expanded circuit that exceeded the original design criterion of P80 160 µm, which was commissioned in the second quarter of 2015.INTRODUCTIONThe New Afton concentrator, located near Kamloops, BC Canada, commenced operation in June 2012. The mill achieved the design throughput of 11,000 tonnes per day average during the month of September 2012 and, following various operational optimization efforts (O’Hara et al, 2015), achieved the design p80 grind size of 160 µm at the design milling rate in July 2013.During August 2013, trials were carried out to determine the throughput rate limits of the mill circuits and to evaluate mill metallurgical performance at higher rates. Milling rates up to 15,500 dmt per 24 hours were achieved during this trial, demonstrating potential capability for sustained milling rates exceeding the nominal design capacity. As was expected, the higher milling rates resulted in a coarsening of the grinding circuit product size, and resulted in marginally lower recoveries in the rougher flotation circuit. Despite the lower recoveries, the higher milling rates resulted in higher metal production and the financial analysis favored running at rates of at least 13,000 tonnes per operating day, with the higher cash flow in the near term offsetting the value of metal production loss at the end of the mine life (due to the short term increase in the rate of reserve depletion). Milling rates were sustained above 13,000 tonnes per day through the second half of 2013, and increased above 14,000 tonnes per day by mid-2014 (Figure 1)."

Jan 1, 2016