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By Andrea McGlynn

Iron Ore Company of Canada, Labrador City 2011?Canada's largest iron ore producer, leading global supplier of iron ore pellets and concentrates ?Rio Tinto (58.7%), Mitsubishi Corporation (26.2%), Labrador Iron Ore Royalty Income Fund (15.1%)

May 1, 2011

By J. E. Nesset, S. R. Rao, J. A. Finch

"For base metal sulphides, iron rejection starts in mineral processing. This review focuses on changes in plant practice specifically to improve iron sulphide rejection by control of contaminant ion effects and discusses selected general innovations in mineral processing that can be expected to play a role. One group of changes to control contaminant ions is in size reduction (the use of stage grinding and inert stirred milling) and another is in system chemistry (order of reagent addition and exploitation of the flexibility of sulphoxy reagents). Plant examples illustrate the successes, indicating that there are several options available. The general innovations start by considering the new tools for measuring gas dispersion properties and quantifying froth characteristics by imaging. Examples of their application to improve concentrate quality illustrate the potential. Other innovations stress adapting geostatistics to model variation in milling and flotation response to provide feed forward control and rigorous application of statistical experimental methodology in plant test work. An example from one concentrator shows improvements in concentrate quality from application of basic principles can be impressive.INTRODUCTIONThe more iron control (rejection) that can be accomplished in mineral processing the lower the subsequent metal extraction costs and environmental implications of iron disposal, particularly in hydrometallurgical processes. Iron rejection has been a contributing target to periodic attempts to apply mineral processing to nickel laterites and bauxites but the prime application is control of pyrite and pyrrhotite (iron sulphides) in the treatment of sulphide ores by flotation. In this paper some new concepts and innovations introduced in sulphide flotation plants over the past ten years or so (i.e., since a previous review [1]) will be outlined with examples of how they have contributed to improved “iron control”.To increase rejection of iron minerals, first the recovery mechanism should be understood. There are four principal ones, two physical in origin: 1, via locked particles (i.e., flotation of composite particles due to insufficient size reduction to liberate the minerals) and 2, through entrainment in the water reporting to the float product (an unselective process); and two originating from a response to the system chemistry: 1, true flotation (i.e., iron sulphides becoming hydrophobic) and 2, as a result of iron sulphides forming aggregates with hydrophobic particles (e.g., ”slime” coatings). There has been considerable effort to diagnose which of these dominates iron sulphide recovery; in many cases it is due to “chemistry”, in particular true flotation. The flotation effect often appears related to contamination by metal ions [1]. This can reduce selectivity in two ways: 1, metal ion species may activate iron sulphides, or 2, cause general depression which demands more vigorous flotation conditions (e.g., additional collector) with consequent loss of selectivity. This is where the review will start, by considering ways to control contaminant ions."

Jan 1, 2007

By Paul Krivokuca

Outline Quinsam Coal Mine ?? Location ?? Mine History ?? Safety ?? Production ?? Hiring 4S Mine ?? Depillaring design ?? Structural mapping, Risk assessment ?? Rock Bolt Design Selection ?? Training

May 1, 2008

By W. J. Dick, K. G. Kress

A pipeline syphon can be a simple and reliable alternative for water conveyance as compared to otherwise more complex pumping systems or low-level outlet conduits through earthfill dams. In a simple form, it consists of a pipeline constructed over the top of a dam or other high topographic feature, connecting a source pond to a receiving pond or channel (Fig. 1). The syphon action is initiated by removing air from the pipeline (priming) using water pumps, vacuum trucks, or simple devices that can be purchased “off-the-shelf.” Once primed, the syphon flow rate can be controlled or stopped using conventional gate, valve, orifice plate, or other devices. Repriming is not required even if the flow is stopped, provided the syphon was not drained. A limiting factor in considering a pipeline syphon as a design option is the maximum lift (i.e. vertical distance from source pond water level to summit of pipeline syphon) which is approximately 9 m or less, depending on pipeline syphon features and site conditions. A pipeline syphon can be quickly constructed from readily available materials including steel, concrete, polyethylene, or PVC pipe. It is particularly well suited to many of the unique conditions often encountered at mine sites. This paper identifies situations that can sometimes be best resolved using a pipeline syphon. It provides basic principles for pipeline syphon design and includes case histories of their successful application.

Jan 1, 2005

By J. Goyman, K. R. Rawling

"1.1 Location and Transportation ServicesNanisivik Mines Ltd. is located at latitude 73°, 3,200 km north of Montreal and 700 km north of the Arctic circle, on the south side of Strathcona Sound, which is an arm of Admiralty Inlet, in Northern Baffin Island. The orebody and mill site are 4 km distant from Strathcona Sound which provides deep water anchorage for ocean transportation of concentrate to Europe during the short northern shipping season from mid-July to mid-October.The operations and the resident community have direct jet aircraft service from Montreal two days per week, and local nights connect with other Arctic communities in the northern Baffin region. Bulk supplies and general cargo are delivered in the summer shipping season.1.2 HistoryThe Texas Gulf Sulphur Company, now Kidd Creek Mines Limited, staked claims in the Nanisivik area in 1957 and carried out exploration through to 1968. After the Kidd Creek orebody was found, Texas Gulf optioned the property in 1972 to Mineral Resources International Inc. (MRI), which undertook further exploration and completed a feasibility study.Nanisivik Mines Ltd. was formed in 1975, with MRI as the majority shareholder and minority shareholdings held by Metallgesellschaft, Billiton and Texas Gulf. Strathcona Mineral Services Limited was retained to manage the project during both the construction and operational phases."

Jan 1, 1984

By M. Schmidt, P. Nord, M. Larinkari

Cathode stripping machine technology has been developed significantly both in electrorefining (ER) and electrowinning (EW) applications in the last ten years. The main driver has been robotic cathode handling, which has enabled increased flexibility to meet different layout and material handling requirements for both ER and EW processes. Additionally, cathode stripping stations have been continuously developed to optimize speed and to reduce mechanical stress of permanent cathodes. 'taco' type cathodes have been proven as beneficial approach for high capacity operations (particularly in Cu ER), where over 600 cathodes per hour continuous throughput has been achieved. However, the latest developments also apply to split type cathode stripping. Besides material handling, the modern cathode stripping machine has also an important role to provide information for production and quality monitoring. This paper describes the latest developments from the field of copper cathode stripping based on several different projects implemented in the recent years. INTRODUCTION After introduction of the permanent cathode technology in the late 1970s, it became the prevailing technology in copper refineries over the traditional starting sheet process by the end of 1980s. The main driver was less labour-intensive production due to fewer processing steps required along with more automated electrode handling equipment (cathode stripping machine, anode preparation machine, anode scrap washing machine and tankhouse cranes). Since then the design of the cathode stripping machine has followed the development of permanent cathode technology. The main development steps so far can be summarized as follows: Full Deposit Stripping Machine (FDSM) based on waxed permanent cathodes or cathodes with bottom edge strips resulting in split cathodes (copper sheets separated, see Figure 1) Wax less cathode stripping based on V-groove on bottom of the edge strip resulting in either ‘taco’ cathodes (enveloped copper sheets, see Figure 1) or split cathodes Full Deposit Robotic Stripping Machines (FDRSM) using 1–5 robots for permanent cathode handling and cathode stacking.

Jan 1, 2019

By Rosaire Emond

Le concept du forage horizontal a été développé voilà de nombreuses années. Différents types de plate-formes permettant le forage horizontal depuis une cheminée ont été conçus par divers manufacturiers ou opérateurs miniers. Entre autres, on connaît bien la griffe de montage Alimak à laquelle peut être fixée une plat-forme sur laquelle est montée une foreuse. En 1992, Ross-Finlay a développé sa propre plate-forme de forage qui à notre opinion améliore la flexibilité du concept du forage horizontal, en augmente la productivité et en diminue le coût d?ensemble tout en assurant une sécurité accrue des travailleurs. Tel que montré à la figure 1, la plate-forme employée à la Mine Kiena est composée d?un toit de protection contre les chutes d?objets et de deux passerelles de forage de forme carrée. Le toit et les passerelles sont fixées à une structure de 6.5 m de longueur équipée de deux assemblages de rouleaux (non montrés à la figure) permettant de faire glisser l?ensemble sur des rails Alimak. L?angle du toit et des passerelles peut être ajusté pour des cheminées dont le pendage varie de 65 degrés à 85 degrés.

Oct 1, 2001

By Steve Wilson, Bern Klein, Jeff Drozdiak, Stefan Nadolski, Andrew Bamber

"In the adaptation of high pressure grinding rolls (HPGRs) to hard-rock mining applications, circuit development has mainly focused on placing HPGRs in the role of tertiary crusher in the preparation of ball mill feed. In order to take advantage of the potential energy benefits associated with HPGR and stirred mill grinding technologies, work was done to investigate the potential of and resultant energy benefits of placing the HPGR in series with a high speed stirred mill in replacement of the conventional cone crusher / ball mill arrangement. The proposed circuit was designed to operate with two stages of HPGR grinding followed by high speed stirred milling.Using a pilot-scale HPGR unit inline with a 20 liter stirred mill, the design criteria, circuit arrangement and consequent comminution energy requirements for a circuit processing copper sulphide ore from Teck Resources’s Mesaba Copper project were determined. These results were then compared to two other circuits, an HPGR / ball mill circuit and a conventional crusher / ball mill circuit processing the same ore.The testwork indicated at a proof-of-concept level that the HPGR / stirred mill circuit could potentially be appropriate for this duty and further work on the concept is merited. While the results indicate that the HPGR component of the circuit was more energy efficient than conventional cone crushing, the energy efficiency of the stirred mill was found to be lower than that of a comparable ball mill undertaking the same duty. However, increased energy efficiency of the stirred mill portion of the circuit is possible with further optimization and further work on the concept is indicated."

Jan 1, 2010

By S. Auger, S. McDonald, K. Ward

"This preliminary paper provides a summary of lab and mill flotation investigations into the copper collector at two Canadian operations. The orebodies illustrated in these studies both contain mainly copper and zinc with uneconomic quantities of lead. Silver is also present as a secondary precious metal.Each mill has advanced to the extended mill trial stage with Uniflot SP129 as the sole copper collector. The preliminary metallurgical results have indicated higher copper circuit zinc and lead rejection is possible as shown in Case Study #2. In Case Study #1, improved zinc circuit metallurgy has been realized during all the Uniflot copper circuit collector trials. A slight increase in silver recovery has also been observed in the extended trials at both operations. INTRODUCTIONThe Canadian base metal industry is presently going through a rationalization period in response to low metal prices and low grade ore bodies. The response from flotation mills has been to reduce costs by incorporating new technologies such as energy efficient large cells, low moisture pressure filters and low steel consumption SAG mills. Other optimization strategies have been incorporated to get better metallurgical performance such as improved process control and monitoring techniques.The newer flotation plants have embraced many of the latest equipment innovations but they continue to incorporate off-the shelf reagents like xanthate and methylisobutylcarbinol (MIBC) . New collector chemistries have been mostly unsuccessful. Our approach incorporates the best of the old and new chemistry into proprietary formulations that match the application. We have found most single chemical formulations cannot meet the demands of the optimized modern flotation plant."

Jan 1, 1992

By S. Kelebek, E. Yalcin

"Two composite samples of a gold-bearing ore from North Western Quebec (Barry Township) were tested to assess their grindability, recovery by flotation and Knelson concentrator as well as extraction using cyanidation.Work index values (BWi) for two composites were found to be 10.4 and 11.0 kWh/tonne, respectively. Flotation tests involved feed samples for which the particle size ranged from a P80 value of 205 ?m to 53 ?m, which resulted in gold recoveries of 91.8% to 95.8% (at gold grades of 130 to 140 g/tonne). Cyanidation tests indicated gold extraction levels ranging from 94.2 to 97.5%. Higher, recoveries were obtained at finer grind sizes and higher cyanide strengths. However, the grind size had a bigger impact on gold leaching. Gold grade and recoveries obtained with the Knelson concentrator device were lower than those obtained by flotation. Mass recoveries to concentrate products strongly depended on the fluidization pressure used. High gold recoveries can be obtained generally at low water pressures of 3 to 4 psi. Higher water pressures produced more selectivity at the expense of higher gold losses.It can be concluded that the gold ore samples tested have good characteristics due to moderate levels of energy requirement, high gold recoveries and extraction rates attainable.INTRODUCTIONBench scale testing of ore samples is an important step in the long series of metallurgical and economic studies involved in bringing new ore bodies into production (Coleman, 1980). Through a laboratory testing program, the behaviour of an ore in a particular process on a larger scale can be established. In an exploratory manner, this testing can be helpful for classification of ores in terms of “good ore” or “problematic ore” by providing preliminary information on recoverability of target minerals. Potential problems if any can be assessed by further investigation of products (e.g., middlings) obtained as a result of testing. A detailed bench scale experimental program in the metallurgical laboratory can also provide the background work for flowsheet development. Depending on the simplicity or complexity of the ore to be tested and type of details required from a testing program, the amount of samples required in the mineral processing laboratory can vary from several kg to more than one hundred kg depending on how detailed the metallurgical testing program is."

Jan 1, 2008

By Weng Fu, James Vaughan, Stefan Lakemond

Uranium bearing copper sulphide flotation concentrates produced from iron oxide copper gold uranium (IOCG–U) ore contains radionuclides such as 238U, 230Th, 210Pb, 226Ra, 210Po, which have the potential to reduce their marketability, thus posing a challenge to the industry. Thus, a two–stage leach process has been proposed in this study. In first stage, a sulphuric acid leach process effectively dissolve most of the 238U and 230Th into solution, but other radionulclides (mainly 210Pb, 226Ra and 210Po) remain associated with copper sulphide minerals owing to their low solubilities in sulphuric acid solution. In the second leaching stage, the use of methanesulfonic acid (MSA, CH3SO3H) is investigated to selectively dissolve the 210Pb, 226Ra and 210Po from the residue of the first stage. The results indicate that 96% of 238U, 87% of 230Th, 90% of 210Pb, 86% of 226Ra and 17% of 210Po are dissolved using 3 mol/L MSA solution at 90°C for 3 hours, which represents a dramatic decrease in the radioactivity of the copper concentrate. INTRODUCTION Australia is host to one of the largest iron oxide copper–gold–uranium (IOCG–U) deposits in the world. Most of the value in these deposits is associated with copper. In mineral processing, the gangue minerals are separated from copper sulphide minerals that report to a copper flotation concentrate. However, due to the strong association of uranium with copper minerals and hematite, even clean copper concentrates can contain significant amounts of uranium (238U) and its radioactive decay elements, of which thorium (230Th), radium (226Ra), lead (210Pb), and polonium (210Po) are the long life radionuclides. These hazardous radionuclides negatively affect the marketability of the copper concentrates. Customer copper smelters in China, Japan and Europe may consider the radionuclides penalty elements or may not be able to process the feed. Therefore, there is a strong motivation to remove radionuclides from copper concentrates at the mine site before shipping to custom smelters in global market. Hydrometallurgical leaching has been widely employed to remove radionuclides from different kinds of minerals such as copper concentrates (Ragozzini & Sparrow, 1987), zircon concentrates (Aral, Sparrow, McDonald, & Norgate, 2007), and coal ash (Lei, Qi, Sun, Xu, & Wang, 2014). Leaching using sulphuric acid (H2SO4) solutions is the most widely used process for the extraction of uranium from uranium ores due to the relatively low cost and wide availability of the acid (Bhargava et al., 2015; Gilligan & Nikoloski, 2015). The leach process involves treating ground ore slurry with sulphuric acid at elevated temperatures (typically 60–90 °C) under oxidising conditions in order to convert the uranium in the minerals 2+ from the relatively insoluble tetravalent form (UO2) into water–soluble hexavalent state (UO2), which then complexes with SO42- and forms UO2(SO4)22- in solution. The uranium oxidation is generally carried out in the presence of ferric ion (Fe3+), regardless of the reagent and oxidant used. The most commonly used oxidants for uranium include sodium chlorate, pyrolusite and oxygen (Ring, 1980). These oxidants act indirectly by converting ferrous sulphate, formed by the oxidation reaction of UO2 in slurries, into ferric sulphate. The presence of Fe–containing gangue minerals such as hematite in the ore usually provides sufficient Fe required for the oxidation of UO2. After sulphuric acid leaching, the locked thorium (230Th), radium (226Ra), lead (210Pb), and polonium (210Po) elements are chemically liberated from hematite and uranium mineral grains (Ciobanu, Wade, Cook, Schmidt Mumm, & Giles, 2013). However, the extraction rate of radium (226Ra), lead (210Pb), and polonium (210Po) in sulphate solution is very low due to the limited solubilities of (Ra, Pb, Po)SO4 (Bagnall & Freeman, 1956; Liu & Papangelakis, 2006; Nikitin & Tolmachev, 1933).

Jan 1, 2019

By Samantha Beeson, Collin Poole

"This paper examines the development of two new selective flocculation techniques for recovering KCI values from low grade carnallite potash ore.Direct recovery of KCI was achieved using one of two mid-range molecular weight polymers, one non-ionic and the other strongly cationic, in conjunction with the same selective anionic dispersant. From initial response studies on pure minerals and selective flocculation tests on synthetic mixtures, investigations were extended to recover KCI values from a natural carnallite potash ore. A single-stage process achieved products grading up to 61.3% KCI at recoveries up to 85.4% for the cationic flocculant suite and up to 64.8% KCI with recoveries as high as 90.3% for the non-ionic flocculant suite, with dispersant and flocculant dosages of 150-250gr1 and 10-75gr1, respectively. These results clearly demonstrate not only the feasibility of treating such problematic ores at low reagent dosages, but also the potential elimination of desliming prior to froth flotation.The experimental results presented are considered in terms of the reagent adsorption phenomena present in the system and various aspects of floe formation and cleaning.IntroductionAs high-grade sylvinite potash ores become scarce, lower grade sylvinite and carnallite potash ores are becoming more attractive. However, these ores often contain large amounts (up to 15% by weight) of water-insoluble slimes which must be removed prior to KC! flotation, owing to their high adsorptive capacity for the amine collectors used. The slimes are generally removed by scrubbing followed by hydroclassification and any residual material not eliminated is blinded with reagents such as guar gum or starch to inhibit interference in subsequent KC1 flotation.An alternative approach used commercially in Canada by Agrium Inc. involves water-insoluble slimes flotation prior to KC! flotation. In this process, the slimes are selectively flocculated and then separated by froth flotation. Losses of KC! to the slimes product and associated brine requirements are both reported to be less than those encountered in conventional desliming, but the technique is somewhat limited in its application at present."

Jan 1, 1997

By P. J. Mehrfert

"Recovery of mineralized coarse particles is notoriously low in conventional concentrators that treat copper porphyry ores. The application of coarse particle flotation techniques has the potential to achieve comparable metallurgical performance at significantly coarser primary grind sizings. ALS Metallurgy has been conducting both internal research and commercial test programs to investigate the potential of Eriez HydroFloatTM technology on copper sulphide ores. Test data indicates that significant improvements in the recovery of copper sulphide bearing particles between 200µm to 850µm can be achieved using Hydrofloat technology compared to conventional flotation techniques. This paper summarizes the results from a selection of test programs that includes recovery by size data from both conventional flotation and HydroFloatTM tests. An explanation of the apparatus and test procedure will be given. Mineralogical analyses of the test products and characteristics that limit metallurgical performance will be discussed.Description of Apparatus and Test ProtocolThe HydroFloatTM laboratory cell is a clear plastic column with an internal diameter of 14 cm. Water and air are introduced to the cell through a circular manifold located in the lower portion of the cell. The cell has a conical bottom which directs slurry to a pinch valve that controls the discharge of tailings. Feed solids are introduced to the upper portion of the cell in slurry form. The cell is operated in a manner such that the manifold water creates a teeter bed zone of suspended solids in the middle portion of the cell. Tailings exit the bottom of the cell at a high pulp density, and most of the introduced water overflows the top of the column and is collected in a launder. Particles with a hydrophobic surface may attach to air bubbles that percolate through the teeter bed, rendering them sufficiently buoyant to be transported to the overflow launder.Three significant features of this arrangement enhance the potential to recover coarse, low grade particles by flotation techniques. Firstly, the hindered settling conditions created by the rising column of water allow coarse particles to become buoyant once attached to a modest sized air bubble. Secondly, the flow conditions within the cell are relatively quiescent compared to a conventional flotation cell that utilizes an agitator. This allows particles that have limited hydrophobic portions of their surface, due to a low content of exposed sulphide mineral, to maintain a somewhat fragile attachment to air bubbles long enough for successful recovery to the overflow. Finally, the absence of a froth phase at the top of the cell allows particles attached to air bubbles to immediately exit the vessel, reducing the potential to drop back into the pulp phase as bubbles burst or coalesce."

Jan 1, 2017

By John G. Payne

The Pebble deposit contains three contiguous zones. In the Pebble West zone, granodiorite stocks intrude gently warped volcanosedimentary rocks that had been intruded by diorite sills and two felsic intrusion breccias. In the Central zone, volcanosedimentary rocks are cut by diorite and granodiorite sills; the latter are genetically associated with the granodiorite stocks. In these zones, widespread potassic alteration consists of brown biotite and K-feldspar, with abundant quartz-carbonate-sulphide vein stockworks. Chalcopyrite/pyrite ratios range from 2/1 to 1/20. Diorite hosts the highest-grade mineralization, with much of it being disseminated. Patchy, later phyllic alteration is moderate and commonly depleted in copper. The Pebble East zone occupies a granodiorite stock and intruded host rocks below an eastwardly thickening wedge of barren Tertiary volcanic and sedimentary rocks up to 600 m thick. Brown biotite alteration and copper mineralization are strong, with chalcopyrite/pyrite ratios commonly over 1/1. The zone extends to a depth of at least 1200 m below surface; some drill holes bottomed in strong mineralization. Later, weak to moderate sericitic alteration removed some chalcopyrite and locally remobilized gold. The Pebble deposit has a Measured/Indicated resource of 3 billion tonnes grading 0.28% Cu, 0.32 g/t Au and 0.015% Mo (= 0.5% copper-equivalent) and an Inferred resource of 1.13 billion tonnes grading 0.30 g/t Au, 0.27% Cu and 0.015% Mo (0.55% copper-equivalent). The Pebble East zone has an Inferred resource of 2.13 billion tonnes grading 0.55% Cu, 0.34 g/t Au, and 0.037% Mo (0.97% copper-equivalent). The Pebble East zone is open to the east and to a lesser extent to the north, south, and to depth.

May 1, 2006

By K. Loutet, A. Guenkel, K. Nikolaisen

"The INCO SO2/air process was patented by INCO in 1985. Well over 80 mining operations worldwide are using this process to oxidize cyanides to cyanates in the tailings from gold, silver and copper mining operations. The sulphur dioxide (SO2) for this process can be supplied either as liquid SO2, SO2 gas or as a sulphite-based salt. The majority of SO2/air operations have used sulphite-based salts as the ‘SO2’ equivalent probably due to low capital cost. Liquid SO2 has also been used frequently, but due to concerns with safety and availability, this option is often not preferred. The solid salt sodium meta-bisulphite (SMBS) is commonly employed, but its cost is an order of magnitude higher than that of sulphur. Due to the high cost of SMBS, the increased capital cost for a sulphur-burning SO2 gas plant (green field or retrofit) usually has a pay-back of less than three years. In addition, a SO2 gas plant offers potential for significant heat recovery from the sulphur burning process. In terms of safety, the SO2 gas plant only requires the storage of relatively benign sulphur and the site inventory of SO2 will be limited to the plant gas volume. Therefore, a SO2 gas plant will not trigger the Canadian Environmental Protection Act (CEPA) E2 Regulations or the US Clean Air Act Section 112(r) accident prevention regulations, as may be the case for liquid SO2.Most existing mining operations using ‘SO2’ are gold operations, but in recent years there has been an increased interest in larger sulphur dioxide gas plants for hydrometallurgical facilities that employ leaching of cobalt/manganese minerals or precipitation of manganese/iron (e.g. Tenke Fungurume and El Boleo).This paper focuses on the production of sulphur dioxide gas generated on site and on demand for the use in hydrometallurgical processes. The capital and operational costs of a SO2 gas plant is compared to conventional process alternatives using sulphite-based salts or liquid SO2."

Jan 1, 2016

By Thomas R. Larson

"Many concentrators use re-circulating “plant” water for dilution in their flotation circuits. However, if crusher dust suppressants, ball mill grinding aids, or even agitator bearing grease are introduced into this water, and the water itself is not subject to cleaning steps, then sufficient quantities of these chemicals may appear in the flotation cells and begin to impact the flotation separation. A bench flotation procedure has been coupled with statistical analysis in order to estimate the impact of small quantities of non-flotation chemicals on flotation performance.INTRODUCTIONCleveland Cliffs produces iron oxide pellets for blast furnace feed, and is developing the capability to produce ferrous metallic products to supply a wider customer base. Cliffs manages five North American iron mines, including the Wabush mine sites in Labrador and Quebec. Based on various characteristics of the ore bodies, the flow sheets of the five mines are diverse. For example, Hibbing Taconite uses magnetic separation alone as their primary upgrading method, but Northshore and Empire use magnetic separation followed by flotation, and Tilden uses flotation alone. Last but not least, the Wabush concentrator upgrades primarily with gravity and electrostatic separation techniques.At Northshore, concentrate from the magnetic separator circuit is fed to a reverse flotation circuit. By design, the magnetic separators operate at the highest possible recovery of magnetite, as the separator tails are discarded, and relatively high silica content was acceptable to Northshore’s customers. As the steel-making industry moved toward lower and lower slag volumes, the silica content in the magnetic separator concentrate became a problem. The flotation step was added in the late 1980s in order to decrease the silica assay of the magnetic separator concentrate to meet the current customer requirements. In this setting, the flotation feed stream contains very little silica that occurs as liberated quartz particles. As the degree of liberation of the flotation feed stream is then a limiting factor for silica removal, the online grind control then was linked to flotation in the overall plant control system so that grind can be adjusted to achieve improved liberation if necessary for the flotation process."

Jan 1, 2002

By M. J. McLeod, D. R. Lux, R. P. Taylor

"The Mount Douglas Granite constitutes the eastern part of the Saint George Batholith and contains the youngest, most differentiated phases of the batholith. It also contains extensive endogranitic, low-grade. Sn- W-Mo-bearing sheeted vein systems. Early phases of the Mount Douglas Granite are coarse- to medium-grained, seriate granites that represent the slowly cooled carapace of the intrusion. Late phases are medium- to fine-grained, equigranular and porphyritic granites that were accompanied by repeated and widespread fracturing of the carapace and injection of mineralizing fluids. These granites represent a highly differentiated suite with ""A "" type affinities developed in a ""within plate"" transitional to ""syn-collision"" tectonic setting. 4oAr/ 39Ar dates for biotites from the various phases indicate emplacement circa 367 Ma.The linear distribution of some phases in the Mount Douglas Granite and the systematic orientations of sheeted veins along tensional fractures indicate structural control on granite emplacement and vein distribution. This type of relatively passive endogranitic mineralization is unique in the granites of southern New Brunswick and contrasts markedly with highly fluidized systems operative elsewhere in the area. IntroductionInvestigation of Sn-W-Mo occurrences in New Brunswick has shown that most of this mineralization is associated with late, highly differentiated phases of post-orogenic (Acadian) granitoid magmatism that occurred during Devonian and Early Carboniferous time (Ruitenberg and Fyffe, 1982; Taylor et al., 1988). The polymetallic Mount Pleasant Sn-W-Mo deposit (Ruitenberg, 1967, 1972; Kooiman et al., 1986)and the Burnt Hill W deposit are two orebodies of this type in New Brunswick. Numerous smaller deposits associated with these granitoid rocks in southwestern New Brunswick have been documented in previous investigations (Ruitenberg, 1967, 1968; Ruitenberg and Fyffe, 1982; Taylor et al., 1985; Lentz, 1985)."

Jan 1, 1988

By Colin Hardie, Simon Pelletier, Rakan Rhabani, Guy Comeau

"In order to improve the nickel concentrate thickener operation at Raglan Mine, two pulp deaeration cyclones were installed. These units obtained from Outotec Pty Ltd, also called the FrothBuster®, have been in continuous operation since mid-October 2006. The primary objective of the de-aeration cyclones is to lower the solids content in the thickener overflow by de-aerating the concentrate pulp before it enters the concentrate thickener. This dearated slurry should help decrease the nickel particle losses that report to the thickener overflow because of the flotation effect within the thickener.Optimizing the de-aeration cyclones has been a slow and ongoing process, however, initial testing is encouraging. The latest test results indicate a 50 % reduction (0.070 to 0.035 % solids) in nickel losses to the concentrate thickener overflow resulting in an additional $350,000 in annual revenues.This paper will present an overview of installing the de-aeration cyclones within a limited space, as well as the preliminary commissioning and testing challenges. This paper will also discuss the most recent concentrate thickener operating results since the de-aeration units were installed.INTRODUCTIONThe Xstrata Nickel Raglan Concentrator, located at the northern limit of Quebec’s Nunavik Region has been in production since 1997. In early 2002, the concentrator implemented the Zero Process Water Discharge (ZPWD) project to recycle the process water to decrease fresh water consumption and reduce discharges to the environment. Due to this initiative, the chemistry of the process water changed significantly resulting in flotation concentrate streams, which were very frothy and difficult to manage. This phenomenon has resulted in operational challenges in many areas of the concentrator including slurry pumping, flow measurement and concentrate thickening.In October 2005, the Raglan Mine underwent a conversion from an AG mill to a SAG mill operation in order to increase throughput and to optimize key process areas. As part of this project, the first two North American pulp de-aeration (FrothBuster®) units from Outotec were installed to improve the operation of the concentrate thickener. The primary objective of the deaeration technology was to lower the solids content in the thickener overflow by de-aerating the concentrate thickener feed pulp. This in turn was to reduce the froth, which accumulates on the thickener surface, and more importantly reduce valuable nickel losses to the thickener overflow because of the flotation effect within the thickener."

Jan 1, 2008

By Graham Holmes, John Starkey

"The Kinross Kubaka Project is a gold mine located in the Russian Far East. The grinding circuit, comprised of a 6.1 m diameter x 2.74 m long SAG mill and a 4.1 m diameter x 5.5 m long secondary ball mill in closed circuit with cyclones, was built using SAG Power Index and Bond Ball Mill Work Index testing to determine the grinding energy requirements. Extremely cold climatic conditions and the need to keep heated space and capital costs to a minimum, dictated a grinding circuit with no in-circuit crusher and no conveyors external to the mill. High energy requirements for both the SAG mill and the ball mill were correctly predicted and have allowed grinding production tonnage forecasts to be met. Design test data and production results are given.INTRODUCTIONIn 1995 Davy International Canada, now Kvaerner Metals E & C, was commissioned as the Engineer and Project Manager for the Omolon Gold Mining Company, to design and construct a gold extraction plant for the Kubaka gold deposit, located in the Russian Far East. This deposit had been studied extensively in Russia during the exploration phase and a Feasibility Study was completed in 1993 by Kilborn who visited the site twice during that year to inspect facilities and to take metallurgical test samples. More samples were taken in 1995 to check ore hardness variability, as part of the final design program by Davy.The client’s objective in designing the Kubaka grinding circuit was to achieve design tonnage at all times, even when milling the hardest ore zones in the deposit. Since the samples taken in 1993 were taken as composite samples, it was necessary to take additional samples in 1995 to determine the extent of the variability. Six samples were sent to Canada for SAG Power Index (SPI) testing. Five of these were combined as a composite open pit sample, while the sixth sample was taken from the site adits as a composite sample to represent undergound ore. These two composite samples were tested to determine the MacPherson Autogenous Work Index for comparison with the SPI values."

Jan 1, 2001

By M. Courchesne, C. Barbery, D. Hodouin, C. Carrier

"The carbon-in-pulp process for recovery of gold dissolved in dilute cyanide solutions, in the presence of solids, has now replaced the convedntional processes of pregnant liquid/barren solids separation followed by Merrill-Crowe precipitation. Most plants built over the past ten years incorporate a CIP section. The process is described as cheaper in capital and operating costs, and few major start-up or operational difficulties are reported. For the purpose of the present paper, the CIP process will be assumed to consist of the following sequence of operations (Figure 1):- the slurry stream coming out of the cyanidation circuit is first screened to remove any tramp oversize material. Gold dissolution is assumed to be complete before entering the CIP circuit.- the slurry is sent into a series of agitated tanks in which it is mixed with activated carbon. The carbon particles size is much coarser than the top size of the ground ore. The tanks are fitted with screens that allows the slurry to overflow continuously into the next tank, while retaining the carbon in the tank. Slurry dilution is adjusted to prevent carbon from settling in the tanks. Agitation intensity is kept low in order to minimize carbon abrasion.- carbon is transfered from tank to tank in a discontinuous manner. According to a specific schedule, the carbon from tank one is removed, then carbon from tank two is sent to tank one, and so on, till the last tank, which is filled with regenerated carbon corning from the elutionregeneration section. The amount of carbon removed, the transfer from tank to tank, the sequence of transfer, the carbon inventory in the various tanks are parameters which are, up to a point, at th2 discretion of the operators, as long as they remove, batch wise, an amount of loaded carbon containing a similar quantity of gold to the one that has entered the circuit since the last transfer."

Jan 1, 1987