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By P. O. Rag, Daniel E. Walsh

"Gravity concentrators are evaluated either in closed or open circuit. The process developed by MIRL employs the use of neutron activated gold (/98Au) and radiation detection circuits which monitor the concentrator's product pulp streams. Concentrator recovery is determined from the ratio of the number of gold particles detected as passing into the concentrate to the total number of gold particles detected over all monitored pulp streams. This system eliminates the experimental errors introduced by sampling, sample reduction, and assaying and allows recovery to be evaluated for gold particles of specific sizes, shapes, and flatness. In addition to allowing for a more accurate and detailed recovery analysis, this process significantly reduces the time and cost required for unit evaluation when compared to other gold recovery analysis procedures. This system enjoys its greatest utility where a large number of tests are to be run in order to evaluate a concentrator over a variety of operational settings, i.e. factorial designs, and allows accurate testing of concentrators operated in closed circuit systems.IntroductionThe surge in gold prices of the late 1970s and early 1980s was followed closely by a renewed interest in gravity concentration processes, especially in concentrators claiming efficient recovery of fine placer gold. Cones, spirals, spinning bowls, jigs and cyclones are all considered potential fine gold concentrators. Fine gold, in this context, is defined as not only that gold finer than 250 um (60 mesh""), but also coarser gold, whose low shape factor may decrease its recovery by conventional, large-scale sluicing operations.Adoption of fine gold concentrators by the Alaskan placer milling industry was retarded by the lack of reliable, accurate, and reproducible data on concentrator efficiency with respect to fine placer gold recovery. In most cases, concentrator performance was based on a limited number of tests run on placer material, about which only slight knowledge of gold size and shape characteristics existed. In order to be prepared for the resurgence of interest in fine gold concentrators, which will follow the next upturn in gold prices, a detailed study of the fine gold recovery characteristics of gravity concentrators needs to be conducted."

Jan 1, 1986

By Alireza Sarafi, Alijan Aftabi

"A hydrocyclic system used for mineral separation at the laboratory scale results in the separation of more than 90% of micas. The system functions only by water and might offer advantages where other methods of mica separation are not available. INTRODUCTIONVarious methods for the separation of micas, wollastonite, and garnet, e.g., screening, flotation, magnetic separation, and heavy liquids, are well known (e.g., Spence, 1929; Bates,I960; Gillson et al., 1960; Benjamin, 1971; Connor and Kolhnar, 1991; Collings and Andrews, 1991). However, where these methods are not available, the separation of micas by a hydrocyclic system developed by Sarafi and Aftabi (1991) may prove to be economic. This note is a brief description of the hydrocyclic system. Minerals and MethodFor testing the recovery of mica and other minerals from different rocks (garnet-mica schist, skarn, and pegmatite), standard minerals were tested first, and then rock samples. Pure mineral standards were prepared from coarse-grained samples of pegmatite, mica schist and skarn; the grain size of the minerals separated was more than I cm. The Hydrocyclic System and Mineral SeparationThe hydrocyclic system is shown schematically in Figure 1. In the first step in mineral separation, mineral and rock samples were crushed, milled, and then sieved to minus 250 plus 125 microns. Approximately 10 gm to 25 gm of the sample were inserted in decanter Dt • The motor pwnps, M1 and M2, were then turned on and the water valve rl was opened. At this time, water was flowing turbulently (30 em3/sec. to 50 em3/sec.) and minerals were separated according to their crystallographic nature, density and!or habit. The fInal separation was done by exit valves r2, r3 and r4, while the motor pwnps were still operating. After each experiment, the rt , r2' r3 and r4 valves were closed and the motor pumps were turned off. Following this, minerals in each decanter and the mter were collected, dried and weighed. The whole time of the experiment was optimized to be about 10 minutes, but varied from 3 to 15 minutes for micas and garnets. respectively."

Jan 1, 1994

By Georgy A. Cherkashov, Andrey V. Efimov, Yury S. Borodaev, Nadezhda N. Mozgova, Sergey G. Krasnov, Tamara V. Stepanova, Alexander M. Ashadze

"Abstract - A horizontal mineralogical-geochemical ore zonation with a southeast to northwest trending succession of Cu and Fe (Zn) main zones has been discovered in the Logatchev hydrothermal field (14°45??N Mid-Atlantic Ridge). The major minerals are pyrite, chalcopyrite, bornite, isocubanite, digenite and chalcocite. Zn-sulfide, marcasite and covellite are less abundant. Rare minerals are native copper, zincian copper and cobalt pentlandite (first found in oceanic massive sulfides), Cu and Cu-Co-bearing Fe-disulfides. Two groups of the main sulfide assemblages with different copper contents are recognized: Group I comprises assemblages with primary sulfides (isocubanite, chalcopyrite, pyrite and Zn-sulfide; Cu up to 35 wt% in Cu-bearing minerals); Group II contains assemblages with secondary Cu-rich sulfides (bornite, digenite, chalcocite; Cu ~60 wt%) as well as rare minerals (cobalt pentlandite, native copper and zincian copper). The assemblages of Group I formed at a temperature of about 300°C; the sulfur fugacity (log fs2 atm) at 300°C is probably 9.9. Relations between minerals of this group are contradictory, attributed to non-equilibrium environments, and no unified sequence of crystallization of minerals has been determined. Minerals of Group II replace, rim or cross the assemblages of Group I. Group II minerals crystallized in a distinct sequence (cobalt pentlandite-bornite-digenite- chalcocite-native copper), with increasing copper content from the start to the end of the depositional sequence. This may correspond to pH from 7 to 8 near Eh = -0.2 at 250°C. Minerals of this group were generally formed under hypergene alteration processes. Compositions of ore minerals were studied by means of electron microprobe analysis (EMPA) (150 analyses). Narrow composition fields were revealed in all main sulfides, which comprised known experimental data. ZnS contains 6.0 to 28.0 mol.% FeS, which is close to data known for other hydrothermal massive sulfides from the Mid-Atlantic Ridge (MAR). A common admixture in ZnS is Cu (0.16 to 1.32 wt%). Rare grains of unusual anisotropic Fe-disulfides, (Fe0.95Cu0.05)S2 and Cu(Fe,Co)9S20, were found among the ore minerals."

Jan 1, 1999

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 Cecilia Arrue, Dominic Fragomeni, N. O. Lotter

"A statistically designed benchmark mill survey was carried out in July 2005 at the Kidd Creek Concentrator, D Division circuit, which processes Ni/Cu ore from the Montcalm deposit. A structured statistical survey approach was followed, together with defined quality controls, which would assign a 95 percent level of confidence in the result.A set of six surveys were completed over a limited time; the corresponding stream samples of the surveys found acceptable were combined to obtain high confidence composite samples. The head grade of each survey is compared to two distributions to test its acceptability; surveys that fail to conform to these trials are rejected. The accepted set of surveys, which is a robust representation of the process, is prepared into a composite for QEM-SEM analysis and mass balance purposes.Where the above practice is a standard procedure in Falconbridge for QEM*SEM purposes, in this case each individual two-hour survey unit was measured and analysed for a mass and value balance. It is shown that none of the individual mass and value balances close as well as the composite set of accepted two-hour survey units. This founding is in keeping with the averaging properties of the Central Limit Theorem.INTRODUCTIONMontcalm OperationsNi/Cu bearing ore from the Montcalm deposit is being processed at the Falconbridge, Kidd Creek Concentrator in a modified D Division circuit. The circuit was commissioned in the fall of 2004. Ore is crushed at the minesite to –19 mm and trucked to Kidd for delivey into the 4 parallel fine ore bins.Ore is milled in a rod mill followed by cyclones and 2 parallel ball mills. The grinding circuit treats Montcalm ore at an average feed rate of 100 wet tonnes/hour (750,000 tonnes/annum) and produces a flotation feed particle size of 83% -53 um. A simple flowsheet including bulk Cu/Ni roughing and 2 bulk cleaning stages followed by Cu/Ni separation in 2 stages is practiced. Soda ash is used as a pH modifier in bulk flotation. SIBX is the collector. Lime and Dextrin are added into Cu/Ni separation for the depression of pentlandite."

Jan 1, 2007

By T. Negeri, M. Boisclair

"A hybrid process consisting of flotation and magnetic separation has been developed to concentrate multi-phase rare earth minerals associated with a carbonatite ore that contains a significant amount of niobium. Fine grinding to a d80 of <65 microns was required to obtain acceptable separation of 92% of the contained total rare earth oxide (TREO) and niobium in half the feed mass. Whereas it was possible to obtain 85% of TREO in 28% of the feed mass through staged cleaning, it was demonstrated that at least 87% and 85% recoveries of TREO and Nb respectively could be achieved in 16% of the feed if the rougher-scavenger concentrates were calcined, quenched, reground to <25 µm and then magnetically upgraded. The paper discusses the overall concept of the flowsheet and the experimental strategies that led to this progress. INTRODUCTIONOut of the approximately 250 minerals known to contain rare earth elements (REEs), only about 60 of them contain REEs in significant amounts. Over 95% of the world’s rare earth oxides (REOs) are believed to occur in the form of three minerals (Pradip & Fuerstenau, 1991) namely bastnaesite (Ce,La,Y)CO3F, monazite (Ce,La,Nd,Th)PO4 and xenotime Y(PO)4. An estimated 51.4% of global REO resources are hosted by carbonatite deposits with the same three minerals as major constituents (Weng et. al., 2015). Any one of the REEs or commonly all of them can be found in these minerals (Ravnaas, 2010). As the demand for critical REEs increases due to the ever increasing applications in advanced technologies, many deposits containing other forms of REE minerals will have to be developed. REE minerals are extensively dispersed and intergrown with several oxides, silicates, carbonates, and phosphate minerals. Consequently it is possible to liberate them only at fine grinding and is thus difficult to separate and produce high grade rare earth mineral concentrates. Most of these minerals contain other important elements such as tantalum (Ta), niobium (Nb), uranium (U), and thorium (Th) in varied quantities. REE minerals in the same ore may exist as carbonates, fluorocarbonates, niobates, silicates, phosphates etc. in association with gangue minerals belonging to different mineralogical groups."

Jan 1, 2016

By R. W. Lawrence

"In the past year, BioteQ Environmental Technologies Inc. have initiated design and construction of new plants to add to an existing list of plants that utilize the BioSulphide® and ChemSulphide™ process technologies for metal recovery and water treatment. In these plants, low cost biogenic sulfide or, in some cases, chemical sulfide reagent is used to selectively remove metals from contaminated water. The process produces high grade, marketable, metal sulfide products even from low grade feed solutions. The revenues generated from the sale of products can allow for water treatment at a profit in some cases, offset treatment costs in others, and generally reduce sludge disposal and management costs. The effluents produced from the process can meet stringent environmental standards. Current plants in operation and under construction range in sulphide demand from 50 kg/day to 3.7 tonnes/day, with the feasibility of larger plants with sulfide demands ranging from 6 to 43 tonnes/day under review.Biogenic sulfide reagent can also be considered for other applications in hydrometallurgy and mineral processing. Uses under consideration by a number of mining companies include: (1) replacement of sodium hydrosulphide in the SART process for cyanide regeneration and copper recovery in the processing of gold-copper ores, and (2) replacement of sodium hydrosulphide for use in the separation of copper and molybdenum in flotation. Biogenic sulphide, produced on demand at site, offers significant cost savings and provides the additional benefit of improved safety due to the elimination of the transportation, handling and storage required for chemical sulphide reagents.This paper reviews the applications of the BioSulphide® and ChemSulphide™ technologies for water treatment and metal recovery, and introduces projects where biogenic sulfide can be applied in hydrometallurgy and mineral processing."

Jan 1, 2007

By David deVries, James Rajith, Paul Mascrenhas, Nurzhan Dyussekenov, Lawrence Hanschar

Kansanshi Copper Smelter commenced operation in March 2015. The smelter features four Peirce- Smith Converters (PSCs) operated in a three hot, two blowing configuration. This paper describes some of the process automation and design features implemented to maximize operational and environmental performance. To ensure that the converters were easy to operate, efficient, and were protected in emergency situations such as power failures or loss of tuyere air supply a significant focus was placed on automated sequences initiated by “one button” operator commands or otherwise automatically by the Distributed Control System (DCS). Examples include: main converter roll-in and roll-out sequences; emergency roll-out in the event of power failure; and the charging sequences for reagents (silica and reverts). Other automated features configured for the PSCs were specifically aimed at maximizing converter capacity and reducing the environmental impact from fugitive emissions. These features include: “one button” in-stack charging; automated restriction of blowing rate based on the available acid plant capacity; and fully-automated control logic for the Secondary Ventilation System. In-stack charging allows anode furnace oxide slag to be poured into the PSCs while the converters are still blowing and still sending off-gas to the acid plant. PSC Secondary Ventilation control logic was configured to require minimum input from the operators while ensuring maximum utilization of the available draft from the ventilation system. Highly automated PSC controls contribute to high utilization, averaging 41.2 in-stack hours per day. INTRODUCTION Kansanshi Copper Smelter (KCS) is located near Solwezi at Kansanshi copper mine in Zambia's Northwestern province. The smelter treats concentrate produced at Kansanshi mine as well as Sentinel mine located 150 km west of Solwezi. Four Peirce-Smith Converters (PSCs) are installed at KCS. The nominal capacity of each converter is to treat 350 t of copper matte and produce up to 250 t of blister copper per batch. Normally three PSCs are hot, with two able to operate (produce off-gas), and the third converter on hot standby. The fourth converter is normally under cold repairs. The blister charge size is relatively low for the matte charged due to the unavailability of tank house or other high copper scrap.

Jan 1, 2019

By Tapiwanashe Chenje, David Hewitt, Peter Radziszewski

"Steel media wear is a major cost component in mineral processing. In the case of existing operating mill, steel media wear prediction can be used to help compare the performance of steel media from different suppliers before completing marked ball tests. It can also be used to investigate how ores from different site of a given mine affect steel media wear. In the case of Greenfield cases, steel media wear prediction can be used to estimate the cost of steel media for a given feasibility study.This paper presents a wear testing and prediction methodology that ties together abrasion wear and corrosion wear testing results through the use of a simplified discrete element charge motion model. The experimentation dimension to this methodology takes samples for steel media, mine site ore and water to determine abrasive and corrosive wear parameters of a given media, ore, water set.The test methodology development is illustrated along with a comparison of wear predictions with observed wear rates in some operating mills. A case study is included that illustrates how this methodology was used to estimate steel media wear performance. The paper closes with a discussion on the use of parts of the experimental methodology to investigate other abrasive wear conditions such as liner wear, paste backfill piping and ore breakage as well as avenues for future test development.INTRODUCTIONThe cost of comminution processes represents a significant part of the overall costs of a mining operation. Grinding media/liner wear alone accounts for 15-25% of the total operating cost. The annual steel grinding media consumption is approximately 2.5 mtpa or 0.3% of total world steel production [Smorgon Steel Group, 2005]. Using an estimated grinding media cost of about US$0.75/kg [Halbe and Smolik, 2002], gives a media cost of over US$1.875 billion per annum. Mining continues to shift towards exploitation of lower grade ores as the higher grades get depleted. These low-grade ores require finer grinding for the liberation of the value minerals, and hence it can be assumed that grinding media consumption will also continue to rise.The Bond abrasion test, dating back to 1963, [Weiss, 1985] is still used to estimate the total mill media wear. This is a purely empirical test based on the ore. From Bond abrasion index tests, it is possible to determine a wear rate using the following function:"

Jan 1, 2009

By Derek H. C. Wilton, David T. W. Evans

"Abstract - The Midas Pond gold prospect, central Newfoundland, is hosted by sheared and altered felsic and mafic pyroclastic rocks of the Tulks Hill volcanics, Victoria Lake Group, and has been trenched and tested by 19 diamond-drill holes. Gold values are sporadic, with the best diamond-drill intersection assaying 7.3 g/t over 0.9 m, and the highest grade channel sample containing 14.7 g/t over 1.15 m. Alteration and mineralization are confined to a 200 m wide brittle-ductile shear zone. This shear zone formed in response to regional D1 deformation, and the shear zone fabric parallels the regional S1 foliation. D2 deformation resulted in broad Z-shaped flexuring of the shear zone. The deformation was probably related to the ca. 396 Ma to 420 Ma Salinic Orogeny.Advanced argillic alteration and an extensive halo of Fe-carbonate and pyrite surround the gold mineralization. Alteration minerals include pyrophyllite, paragonite, quartz, plagioclase, chlorite, fluorite, Fe-carbonate, and pyrite. As such, the alteration seems to reflect combined epithermal (argillic) and mesothermal (carbonate) mineralizing environments. The argillic alteration and fluorite represent a regional hydrothermal effect on the felsic volcanic rocks; whereas, the Fe-carbonate and pyrite represent a mesothermal alteration associated with auriferous quartz veins. The Au, which is spatially associated with pyrite, occurs in three structurally-controlled vein sets. These veins are confined to the contact between a highly deformed breccia (the banded mafic tuff) and structurally overlying, altered felsic tuffaceous rocks. The vein sets include: V1 boudinage veins that parallel the shear-zone fabric (C-shear veins), and V2 and V3 extensional fracture veins which are controlled by an S2 fracture cleavage. V1 veins are the earliest and contain the lowest concentrations of Au. V2 and V3 veins are concentrated within the hinges of the D2 flexures.Pb isotope data for galena separates indicate that the Pb in the vein systems was probably of local derivation. Sulfur, C, and O isotope data for mineral separates are similar to data for typical mesothermal lode gold deposits. Fluid inclusion data suggest the presence of at least three hydrothermal fluid types in the region, viz., (1) an early saline fluid which produced the argillic alteration, (2) a CO2-bearing fluid directly related to the auriferous mineralization, and (3) H2O-NaCl fluids which flowed through the Midas Pond shear zone at a lateral distance from the zone of mineralization. The mineralizing fluids are interpreted to have been CO3 -2-rich types which originated through metamorphic dehydration and decarbonation of mixed island-arc and continental crustal rocks. Gold precipitation resulted from the reaction of these fluids with the Fe-rich breccia unit. Pre-existing and simultaneously crystallizing pyrite served as loci for Au precipitation."

Jan 1, 2000

By J. E. Udd

"In early 2000, CANMET-MMSL was engaged by the Ontario Ministry of Labour (OMOL) to conduct an international study of the practices, in several key mining jurisdictions, of determining the criteria that are used to decide when the wire ropes used in mine hoisting are to be retired from service. Subsequently, the author visited key officials and institutions in Canada, the United States, the Republic of South Africa, the United Kingdom, Germany, and Poland. Material from the Australian province of New South Wales and Hungary was also included.The report that resulted included, among other sections: (1) the history of wire ropes and trends in manufacturing technology; (2) a review of mine hoisting practices; (3) the requirements regarding safety factors; (4) a review of the various operations and their potential future development; (5) the criteria for rope replacement; (6) potential future developments that could take place in both hoisting and rope inspection practices in Canada; and (7) recommendations.Subsequently, the report was issued by OMOL, and as a CD-ROM by CANMETMMSL. In this paper the author presents the key findings and the recommendations made in the study."

Jan 1, 2004

By Chris Fleming, John R. Goode, James Brown

"Data will be produced from two uranium feasibility studies that were conducted in 2008, both involving uranium recovery from leached slurry by resin-in-pulp (RIP) with strong base resins. Excellent metallurgical results were produced in RIP pilot plants in both studies, and the loaded resins were eluted with strong sulphuric acid and recycled to RIP in an integrated process. Resin losses by either abrasion or mechanical breakage were immeasurably low.Metallurgical results from these two campaigns will be presented, along with an economic analysis of RIP versus the alternative process of CCD or filtration and solvent extraction.INTRODUCTIONThe recovery of uranium directly from pulp with strong base resins in RIP contactors was practiced commercially in the USA in the 1960s, but the process never established itself as a viable alternative to solid/liquid separation and solvent extraction. This was due mainly to the necessity to do a sand-slime separation partially resulting from the unavailability of coarse resin beads, inadequacies in the design of the RIP contactors and screens to separate resin from pulp, as well as problems with the strength and durability of resins in agitated pulps.There have been significant advances in the last 30 years in each of these areas. All the commercial resin manufacturers are now producing very tough, durable, strong base resins, and are manufacturing them in a large bead size that is better suited to RIP applications. In addition, with the emergence and wide spread implementation of the carbon in pulp process in the gold industry in the 1980s, many innovative techniques for preparing leached slurries and separating carbon from pulp have become firmly established. These techniques have proven to be equally applicable to resin separation from pulps in RIP contactors."

Jan 1, 2010

By L. Parthasarathy, S. K. Pan, M. P. Srivastava

"Total iron ore reserves in India, as reported by the Indian Bureau of mines, are 23.6 billion tons. Hematite iron ore reserves constitute about 45 % of the total reserve. If we go by the National Steel Policy-2005 projections, India’s supply of hematite iron ore would not last more than 40 years. Laboratory investigations were carried out for the development of a process flow sheet for treating low grade as well as Banded Hematite Quartzite (BHQ) iron ore. The purpose of these investigations was to aim for better resource utilization.Beneficiation studies were conducted on samples of low grade and siliceous iron ore from the Dalli mines, a captive mine of the Bhilai Steel Plant and BHQ ore from Southern states of India. The process flowsheet, developed for upgrading of low grade iron ore, incorporates wet screening of the ore at 0.5 mm, followed by grinding of the +0.5 mm fraction to below 0.5 mm. The whole –0.5 mm fraction, after grinding, was treated in a conventional hydro-cyclone and stub cyclone followed by treatment in Wet High Intensity Magnetic Separator (WHIMS). This resulted in a concentrate product having >66% Fe content from a feed containing 55.5% Fe with an iron recovery of 48 %. The process flowsheet developed for treating BHQ ore in the laboratory resulted in an increase in iron grade in the concentrate from 43.4 to 61.3 Fe % with a yield of 51.7 % .INTRODUCTIONThe classification of Indian iron ore resources have been made compatible with United Nations Framework Classification (UNFC), which is more scientific and has been adopted in most countries in the world. Iron ore reserves and resources estimated using the UNFC classification system, as of April 01,2000 are about 12,906 million tons of hematite ore and 10,628 million tons of magnetite ore. The Indian Bureau of Mines (IBM) has given a breakdown of these reserves by designation as either lumps or fines of hematite iron ore (Table 1). The percentage of the reserves reported lumps and fines are 42 and 34% of the total reserves respectively. High grade iron ore is less than 10% of the total reserve."

Jan 1, 2008

By Chen Bang, Xia Xiao’ou, Zhang Feng

"Ores with compressive strength lower than 150 MPa are usually easy to be crushed by conventional jaw crushers and cone crushers. However, for hard ores with compressive strength exceeding 150 MPa and 200 MPa and even up to 250 MPa, such as andesite, hard basalt and olivine, it is difficult to reach satisfactory crushing size and yield or difficult to meet service life and stability use requirements of quick-wear parts by using conventional crushers. As continued increase in ore mining depth and constant generation of new type materials, equipment demand for crushing hard ores becomes increasing high. A new type inertia vibrating crusher is designed. It is powered by a high-speed rotating vibration exciter and can obtain required crushing force by adjusting eccentric mass, eccentricity and rotating speed so as to crush various ores with high hardness. This paper introduces the design process of its structure and working principle, discusses the design calculation methods of vibration excitation system, vibration damping and energy storage system and other key mechanism and provides technical references for this type of vibrating crusher.DESIGN PHILOSOPHYThe conventional jaw crusher mainly acquires extrusion force from crank motion to realize crushing. The motor drives the eccentric shaft via the transmission device to bring the moving jaw to swing relative to the stationary jaw. When the moving jaw moves close to the stationary jaw, ores between them are crushed by extrusion; when the moving jaw moves away from the stationary jaw, crushed ores are discharged from the discharge gate at the lower end between the moving jaw and the stationary jaw.The conventional spring cone crusher and the hydraulic cone crusher mainly acquire extrusion force from cranking motion to realize crushing. The motor drives the eccentric sleeve via the transmission device, the moving cone shaft generates an eccentric distance to make cranking motion under the action of the eccentric sleeve. When the moving cone moves close to the stationary cone, ores between them are crushed by extrusion. When the moving cone moves away from the stationary jaw, crushed ores are discharged from the discharge gate at the lower end between the moving cone and the stationary cone under the action of dead weight."

Jan 1, 2016

By Trevor Yeomans

"NVI Mining’s Myra Falls Operation is a forty year old underground mining operation feeding a 4000 mtpd concentrator.The current Mill, built in 1985, was designed to process a large massive sulphide ore-body. This ore-body, known as the “HW”, contains about 50% pyrite. Value minerals are chalcopyrite, and sphalerite, with precious metals occurring in all mineral types.By 1997, mining of a new ore-body, known as “Battle-Gap” commenced. Mineralogically complex, the new orebody contains about 15% pyrite, copper minerals including primary chalcopyrite, secondary bornite, chalcocite and tennantite, combined with increased galena and sphalerite.Flotation performance generally decreased with increasing amounts of Battle-Gap ore types, with both copper concentrate quality (increased lead and zinc content) and copper recoveries being affected.Following significant metallurgical testwork, a new fully sequential copper-lead-zinc flotation flowsheet and reagent scheme was developed.This sequential flowsheet was progressively installed and commissioned during 2006.Copper concentrate lead and zinc smelter penalty contents have been reduced by some 40%, whilst increasing copper recovery by 10%.INTRODUCTIONNVI Mining operates the Myra Falls Operation near Campbell River on Vancouver Island, BC. The complex has been in operation since 1966 with mining being predominantly underground. One unique feature of Myra Falls is that the mine operates within a BC Provincial Park.In 1985, a new large massive sulphide ore-body (HW) had been defined, and metallurgically tested. A new concentrator was built based upon this massive pyrite, chalcopyrite, and sphalerite mineralization producing copper and zinc concentrates by conventional flotation methods."

Jan 1, 2008

By Chuck Edwards

"Cameco operates the high-grade McArthur River uranium mine in northern Saskatchewan, and is constructing the high-grade Cigar Lake uranium mine. Cameco has studied the concept of underground milling of high-grade uranium ore. Compared to conventional milling on surface, the underground milling scheme appears to offer significant cost savings and a lower environmental impact. This paper describes the underground milling scheme, presents process flowsheets and plant layouts, and provides an assessment of potential benefits from underground milling.INTRODUCTIONCameco CorporationCameco is the world’s largest uranium producer, with three operations in Saskatchewan (Rabbit Lake mine/mill, McArthur River mine, and Key Lake mill), two in the United States (Crow Butte and Smith/Highland ISL operations) and two new mines being developed in Canada (Cigar Lake project) and Kazakhstan (Inkai project). Cameco provides uranium fuel refining and conversion services at its Blind River and Port Hope facilities in Ontario. Cameco is part owner (31.6%) of Bruce Power that leases eight nuclear power plants located in southern Ontario. Cameco is the majority shareholder in Centerra Gold Inc., which operates gold mines in the Kyrgyz Republic and in Mongolia.Global Uranium Production RequirementsCentral to Cameco’s view of future uranium supply requirements is the belief that the world is experiencing a nuclear renaissance. This belief is supported by a number of developments, including:•,?Countries representing half the world’s population are now constructing new nuclear power plants. Many countries without nuclear power have begun to plan for it, seeking reliability, clean air and energy independence.•,?Accelerated growth for this industry is also becoming much more likely. In Asia, primarily in China and India, 92 reactors are under construction, on order or planned to help meet rapidly expanding economic needs. Other Asian countries, such as Vietnam and Indonesia, with large populations and rapidly growing economies, are also expected to follow.•,?In the United States the “Energy Policy Act of 2005”, the first energy legislation introduced in over a decade, was signed. This legislation includes provisions recognizing, for the first time at a national level, the clean air contribution of nuclear energy and its critical role in meeting energy security. Upon signing this landmark legislation President Bush stated, “We will start building nuclear power plants by the end of this decade.”"

Jan 1, 2007

By J. E. Nesset, J. A. Finch, Rosenblum. F., S. Somot

"In a previous presentation at the CMP, evidence was shown for the formation of hydrogen sulphide (H2S) in self-heating experiments on pyrrhotite-rich samples (Somot and Finch, 2006). Oxidation of H2S is strongly exothermic leading to the proposition that H2S is playing a key role in the self-heating process. Subsequently, tests in two weathering apparatus (I and II) were run under varying oxygen levels to further examine conditions leading to self-heating. To characterise their resulting self-heating capacity the samples from the weathering tests were subjected to a modified Stage B test using the self-heating apparatus developed by Rosenblum and co-workers (Rosenblum and Spira, 1995; Rosenblum et al., 2001). Results showed highest self-heating after weathering occurs under low oxygen conditions, supporting the H2S hypothesis. It was observed that highest heating rates are not directly related to highest weight gains of samples.INTRODUCTION“Sulphides” refer to several types of minerals containing sulphur in its lowest oxidation state of - 2. Self-heating is the spontaneous heating of a material without external heat input. It is well established that some sulphides will self-heat on exposure to moisture and oxygen (Farnsworth, 1977; Good, 1977; Rosenblum et al., 2001; Wu and Li, 2005). This process is highly exothermic and if the heat generated is not dissipated, combustion may occur when the temperature reaches the ignition point of the sulphide (typically 350-400 °C). The hazards of sulphide self-heating range from handling hot material, to sulphur dioxide release, to fires. The literature documents self-heating incidents underground, on ships and trucks transporting sulphide concentrates, and in mine waste (tailing) disposal areas (Stachulak, 1990; Lukaszewski, 1969; Headley et al., 1977). Lukaszewski (1969), Farnsworth (1977), and Good (1977) note that sulphide minerals can act as fuel to promote underground mine fires. Based on such findings, sulphide minerals could become classified as hazardous materials, making transport of concentrates and storage of mine waste subject to stricter and more costly regulation. The problem is one the industry is keen to avoid."

Jan 1, 2009

By Takahiro Kubo, Katsunori Yamaguchi, Hidehiro Sekimoto

Equilibrium relation between B2O3 slag, RExOy-B2O3 (RE = rare earths) slag and Fe-C alloy formed by melting treatment of neodymium magnet and B2O3 was investigated. When neodymium magnet was melted together with B2O3 and Fe-C alloy in carbon crucible, 3 liquid phases of B2O3 slag, RExOy-B2O3 slag and molten Fe-C alloy formed and clearly separated by the difference of the specific gravities. In the case that excess amount of Cu was also introduced in that system, molten Cu phase additionally formed. Almost all amount of RE were concentrated in RExOy-B2O3 slag. The concentration of RE in RExOy-B2O3 slag increased with the decrease of the melting temperature. While chromium, phosphorous, nickel, cobalt, and copper were concentrated in molten Fe-C alloy, silicon and aluminum were distributed in B2O3 slag and RE2O3-B2O3 slag.

Jan 1, 2014

By Éric Poulin, Simon Gariépy, Mario Paredes Malca

"While the design of a control strategy depends on the available process knowledge, its implementation depends on the available process information. On-line chemical analyzers are today widely spread in the mineral processing industry. A real-time concentration measurement of chemical components is valuable information that could bring control strategies to whole new levels. However, the use of these measures rarely reaches anything more than process monitoring. The reason is simple: the on-line chemical measurements quality is most of the time insufficient to be directly included in control strategies. The choice is then made to let the critical eye of the engineer or the operator appreciates the measurements before making any decision based on it.Data reconciliation through mass balance is a technique today widely used in the mineral processing industry to increase the accuracy and the reliability of plant data, mainly for accounting and survey analysis purposes. Real-time data reconciliation of on-line measurements appears as a natural coupling of technologies that could bring new process optimization trends. Once overcome the technical difficulties, full valorization of the instrumentation investment could be reached.This paper presents the application of real-time data reconciliation for process control and performance monitoring. It proposes a general approach, discusses challenges and illustrates concepts with industry experiences."

Jan 1, 2006

By Charles S. Ney

"THE WRITER'S acquaintance with Kootenay King mine is limited to a brief geological excursion in the vicinity. This paper has been compiled largely from maps of J. Sullivan, mine geologist in 1952, kindly made available to the writer by Transcontinental Resources; a 1942 private report by W. L. Brown; and personal communications with J. Sullivan. Remarks on regional geology, and the comparison with Sullivan mine, are specifically the writer's.The Kootenay King mine is situated in the westernmost range of the southern Canadian Rockies at an elevation of 7,300 feet. From the Sullivan Mine at Kimberley, B. C., it lies 19 miles due east. Six miles of this distance are represented by the pronounced topographical and structural gap of the Rocky Mountain Trench. Mineral claims were locat-ed in the area prior to 1898. The deposit was first seriously explored in the late 1920s, and a small orebody was then outlined and developed. Mining and milling were undertaken in 1951.52. The orebody, amounting to about 14,000 tons, was mined out, and no significant new ore was revealed.The orebody is a replacement of dolomitic argillite by fine-grained galena, sphalerite, and pyrite. The sulphides are finely laminated in the direction of bedding, following out minor folds faithfully. The grade of ore estimated from diamond drill holes (W. L. Brown, 1942) was 11 ·I per cent zinc, 11 ·I per cent lead, and 3·8 ounces per ton of silver. The mineralized zone extends about 250 feet horizontally, 160 feet vertically, with a width of six to twenty feet.General GeologyThe country rocks are sediments of the Aldridge formation of late Precambrian age. For several thousand feet stratigraphically above and below the mine, the formation consists mainly of thinly laminated argillite and siltstone. Several horizons contain appreciable carbonate. About 2,800 feet stratigraphically below the mine, predominantly dolomitic beds mark the top of the underlying Fort Steele formation. The medium-bedded fine quartzites, which characterize about 10,000 feet of the Middle Aldridge in the Purcell range, have a restricted development amounting to only 2,000 feet in the Rocky Mountains. This member is about 2,200 feet above the mine horizon."

Jan 1, 1949