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By Marcel Vallée

Abstract -Traditionally, samples collected for mineral deposit evaluation have been viewed without explicit reference to reliability, with the illusion that a true value is attainable from a few samples and verifications. For exploration, mineral deposit estimation and production quality control the following definition is proposed: "a sample is a small quantity of material relative to the geological mass it represents, collected according to a systematic procedure, of measurable reliability, from which the grade/quality of the mass which it represents may be estimated, based upon appropriate protocols. This paper will concentrate on the actual rock sampling process. Too often, sample size has been, and is still viewed as a matter of convenience, leaving out representivity considerations, despite the importance of samples on project estimation. Linear samples (chips, channels, drill), panel samples, broken rock or mining samples and very large samples are considered. Collection procedures and quality control protocols to improve the representivity of non-mechanical samples are proposed.

Jan 1, 1998

By Steve Dismuke, Tom Diment

In December 1994, Polaris decided to investigate the use of cemented rockfill (CRF) to supplement the existing frozen rockfill system. Extraction plans for the remaining ore reserves required a backfill with superior stiffness proper-ties and capable of developing high early strengths. Although CRF has been used success-fully at several mining operations, there was the additional challenge of making a reliable cemented product in permafrost conditions. There was also the challenge of procuring and storing one year’s supply (13 000 tonnes) of bulk cement. By the spring of 1995, following three months of research, test work and process design, the decision was made to proceed with the procurement and construction of a 3000 tpd batch plant. An ambitious completion date of December 1995 was set. Team Manufacturing of Langley, British Columbia was awarded the contract, complete with the considerable task of having the batch plant fabricated and ready for sea lift shipment by August, 1995.

Jan 1, 1996

By Mori Fresko, Stuart M. Jones Jr

"This paper is derived from a document that compiles autogenous and semi-autogenous mill installations worldwide, through the end of 2010. Updates have been solicited from all major mill manufacturers and most have responded.A review of AG/SAG mills sold to date and in the last five years indicates that the AG/SAG mill market is very healthy and the trend is larger diameter mills with high diameter-to-length ratios. Ring motors (gearless) are the preferred drives for large mills even though the power transmitted through a gear drive is expected to near 11,000 kW per pinion in the near future.INTRODUCTIONThis paper is derived from a document that compiles autogenous and semi-autogenous mill installations worldwide. The list is maintained by Stuart M. Jones, Jr. (author), and is included in the appendix. This list has been the inspiration for similar papers in the past several SAG Conferences. Updates have been solicited from all major mill manufacturers and most have responded. Information from this list is compiled to illustrate the growth of the AG/SAG mill market and the use by industry and selected geographic areas. It covers the market growth through the end of 2010."

Jan 1, 2012

By J. N. Cram

"The comments given are to help allay generally accepted ideas that the handling of arsenic as arsenic-trioxide (As2o3J in and from off gas pyrometallurgical operation fumes is difficult, hazardous and expensive. It need not be any of these things if the basic and rather unique chemic3l and physical properties are understood and taken care and advantage of through present day technologies. These particular characteristics are:-(1) As2o3 sublimes and re-condenses, at a relatively low temperature;(2) As2o3 is, for practical in-plant purposes, insoluble in water at ambient temperatures and quite soluble in water just below the boiling point;(3) residual suspended and dissolved As 2o3 can be removed from water by complete dissolution with very little sulfuric acid, followed by the addition of small amounts of ferrous sulfate and calcium chlorate, with oxidation to the ferric and arsenite state and precipitation of an insoluble ferric calcium arsenate, thereby lowering the As2o3 content of solutions to acceptable levels. Even lower levels of non-toxic forms of arsenic solids and residual As2o3 in the water solutions can be attained by also using phosphate, and lime, along with the above.A background re past arsenic processing methods, present day operations, plus predictions re future trends, and hazards, are outlined."

Jan 1, 1979

By G. A. Sutton

This paper, describes a case example from San Manuel mine, where in situ copper leaching of the supergene zone was done over a 14 year period (1986–1999) in conjunction with underground and open-pit mining operations of a copper porphyry deposit. In 1993–1994, testing of well fields was done to determine the amount of copper that would be recovered from in situ leaching of the supergene zone, which was composed primarily of chrysocolla and copper-clay minerals. It was determined that the leach fluid would reach 70% (‘sweep efficiency’) of the copper minerals and that 75% of them would be leached for a recovery of 52% of the acid soluble copper. However, the final results from Zone 6, which had the longest leach history, showed that approximately 42% of the acid soluble copper was recovered, indicating that the `sweep efficiency’ was closer to 55%. Comparison of drill-core assay data from two holes, one drilled before (ARM158) and the other after (SMO9607) leaching, also indicate that approximately 42% of the acid soluble copper was recovered. The discrepancy between the predicted (52%) and actual (42%) recovery is probably related to a combination of factors, including difficulty in maintaining fluid saturation, channeling effects and gypsum precipitation. INTRODUCTION In situ copper leaching at the San Manuel mine, Arizona, was initiated in 1986. It was recognized that some of the copper oxide resource was not going to be economically recoverable by conventional open pit and heap leach methods using solvent extraction and electrowinning (SXEW). A high stripping ratio would be required to access part of the resource. At the time, in situ leaching was becoming popular in the uranium industry but the approach was virtually unknown in the copper industry, especially not at a large scale.

Jan 1, 2019

By D. G. Macintyre

Thick westward prograding clastic wedges were deposited along the margin of ancestral North America from Mid-Proterozoic to Mid-Paleozoic time. Periodic rifting and tectonic subsidence resulted in the formation of starved sub-basins along this otherwise passive continental margin. These sub-basins host important sedimentary exhalative barite-zinc-lead-silver deposits. In northeastern British Columbia, a thick succession of Paleozoic basinal facies clastic rocks is preserved within the northwest-trending Kechika Trough. Within the trough, sedimentary exhalative barite and barite-sulfide deposits occur in starved basin sediments of Middle Ordovician, Early Silurian, and Late Devonian age. The last are the most economically sigrrificant and are hosted by carbonaceous cherty argillites and siliceous shales of the Middle to Late Devonian Gunsteel Formation of the Lower Earn Group. The largest deposit, Cirque, has reserves in excess of 35 million tonnes averaging 100J0 (Pb +Zn) and 47 g/t Ag. Other potentially significant deposits include Driftpile Creek, Bear, Mt. Alcock, Fluke, Pie and Elf. Stratiform barite deposits are also common, particularly near the basin-to-shelf transition zone. The Late Devonian barite-sulfide deposits typically are zoned: interlarninated barite-sphalerite-galena-pyrite occurring near suspected vent areas grades outward into bedded barite. Laminated beds of pyrite typically occur in hangingwall siliceous shales, particularly above the inferred vent zone. The apparent absence of epigenetic stringer sulfide zones and footwall alteration implies that the deposits were formed from ponded brines at relatively low temperature. Away from the main deposits, the favourable stratigraphic interval is marked by the presence of thin beds of nodular and laminated barite and thin laminae of pyrite. Thin tuff beds also occur locally in this stratigraphic interval and are evidence for a Late Devonian volcanic event. A paleogeographic reconstruction of the Kechika Trough in Late Devonian to Mississippian time suggests that the barite-sulfide deposits formed in third-order starved basins between uplifted, northwest-trending horst blocks that were capped by shallow-water carbonate banks. The presence of intraformational breccias suggests that fault movement accompanied barite-sulfide precipitation. This tectonic activity may have allowed the escape of heated, overpressured metalliferous brines from permeable reservoirs within the sedimentary pile. Sedimentary exhalative deposits formed where these brines were exhaled as buoyant plumes onto the seafloor, cooled, and ponded in anoxic seafloor depressions.

Jan 1, 1992

Overview of Presentation Why Enter Agreements with First Nations Making it Win-Win ?The Principles The Process Making the Process Work Resource Agmts Defined Resource Agreements are legally binding agreements between a resource company and the First Nation(s) whose Aboriginal or Treaty rights may be impacted by the activities of the proposed project

May 1, 2009

By Sobczyk

The article will present a comparison of classifications of resources according to the JORC Code System with Polish classification, basing on system used in Eastern Europe. The differences between these classifications are significant. The information about mineral resources in Poland is not comparable to the main international system of reporting reserves like JORC, SAMREC, and PERC. Although comparisons between these systems are used, the transfer of particular categories of resources from the Polish system into a JORC category is not simple. Basing on the examples of privatization of Polish coal companies through initial public offering on the Warsaw Stock Exchange, the current practices providing guidelines how to convert resources from the Polish classification to the JORC system are described. Such independent assessment of resources is necessary for development of Mineral Expert?s Report (MER or CPR) which is an integral attachment to the prospectus.

Aug 1, 2013

By A. Jeffkins

Visibility is restricted when operating load haul dump vehicles resulting in a number of serious accidents in the Ontario Mining Industry. One of the leading causes of these accidents is the inability of the equipment operator to see clearly people, objects or hazards around the machine. Previous line of sight testing by the research team found specific machine characteristics were responsible for blind spots and impaired visibility. In order to identify feasible design modifications a computer simulation program, Classic JACK 3.1, was utilized. Program features include human factors analysis, which allows evaluation of line of sight and operator posture in order to decrease risk of musculoskeletal injury. Research within the virtual computer environment consisted of a Visibility Audit and a Target Audit in order to quantitatively determine how certain components of the LHD design affected the line of sight around the vehicle. Results indicate that modifying the LHD light brackets, engine profile, cab posts and other components will enhance visibility. The addition of cameras located on the front and rear of the vehicle were able to enhance the operator?s sight lines to previously obstructed areas. Modifications to seat height and the operators orientation both improved sightlines and decreased risk factors associated with back and neck musculoskeletal disorders.

May 1, 2003

In terms of throughput, product size and power draw, grindcurves provide a measure of the response of the mill performance to changes in mill operating parameters – usually mill filling. Since presenting the principle in 2006, grindcurves have been applied to a wide variety of AG and SAG mills. Mill response has been carefully measured for variations in: mill fillings from 15% to 40%; throughput by a factor of two; power by 50%; feed blends; liner profile; and mill speed from 65% to 80%. The outcomes of this extensive program of work are presented as the measured grindcurves. How they have enhanced our understanding of mill operation and how they may be applied to practical daily mill control is discussed. It is proposed that this technique provides a viable and valuable mill control tool that may find extensive application in the industry.

Jan 1, 2012

By J. W. Keyes

"The original grinding circuit at East Malartic consisted of two 8 by 8 C.A.C. mills in closed circuit with two 72 inch Akins Classifiers and an 8 by 10 secondary mill in closed circuit with two 48 inch duplex classifiers. With this circuit, we experienced a drastic drop in tonnage from one month to another, due to variations in the hardness of the ore. The first step to overcome this difficulty was made by increasing the critical speeds of the mills from 78% to 82%. This resulted in an increase in the secondary mill throughput; but, when the change was made on the primary mills no increase was obtained.The drastic drops in tonnage were still experienced and through working with Allis Chalmers the 8 by 10 mill was converted to a rod mill and the two 8 by 8 n1ills to secondary mills. This change eliminated the two 72 inch Atkins classifiers and the scoops on all the mills, since the 48 inch Akins classifiers sands could then be fed by gravity to the secondary mills through feed pipes.The rod mill and secondary mill discharges were fed to a common pump box using 5 by 5 S.R.L.'s running at 1085 R.P.M. with a twenty-five foot head. When this was done the pumping costs jumped from $105. 00 for 6 months to $1050. 00 in one month, or from 0. 0004 to 2. 75 cents a ton milled. Several factors contributed to this jump in pumping costs as follows:"

Jan 1, 1964

By E. J. Marcotte

"BRIEF DESCRIPTION OF GRINDING CIRCUITCrushed ore at -5/ 16 in. is delivered by a conveyor system from the crushing plant to two 1250 ton fine ore bins. The ore is transferred from each bin by a 30 in. feeder conveyor belt, operated by a Link Belt P.I. V. (variable speed) drive, to a 24 in. conveyor that feeds the 8 x 10 Allis Chalmers rod mill.The rod mill discharges to a common mill discharge pump box. Three 6 x 6 SRL pumps are employed to pump the combined mill discharge pulp to a common distributor box which distributes the feed to two of the three Kreb D lSB cyclones in each of the two banks at a hydraulic pressure of 6 psi.The cyclone overflow at a specific gravity of 1. 25 (i.e. 32% solids) is transferred by gravity to two revolving chip screens for wood chip removal and from there to a pump box. Two 5 x 5 SRL pumps feed the pulp to the thickeners. The underflow from the cyclones is fed to two 8 x 8 Allis Chalmers ball mills which discharge to the common mill discharge pump box. Thus, a closed circuit is formed between the ball mills and their respective bank of cyclones. The circulating load as most recently calculated is 262%."

Jan 1, 1966

By Shamil Shaikhutdinov

CO adsorption on gold particles deposited on well-ordered alumina and various iron oxide films has been studied with temperature programmed desorption (TPD) and infrared reflection adsorption spectroscopy (IRAS). Structural characterization has been performed with scanning tunneling microscopy. Data show that the CO adsorption exhibits a size effect in that small particles adsorb CO more strongly. The effect generally vanishes after annealing to 500 K. For annealed samples, CO TPD and IRAS spectra are nearly independent of the particle size and supports studied. The results indicate that the CO interaction with gold surfaces is essentially similar and involves low coordinated atoms.

Oct 1, 2003

By George Hope, Dave Green, Jacques McMullen

The Est Malartic mill has been treating massive sulphide, copper/gold ore from the Bousquet 2 mine since 1989. The need to minimize copper content in flotation tails which are leached for gold recovery and to produce a saleable copper concentrate led to the evaluation of Conventional, Column and HG flotation methods. The HG method was preferred and two full-scale cells were installed and operated over a two-year span. In 1992, it was decided to expand the mill capacity from 1500 tpd to 2500 tpd. A total of nine additional HG Tank Cells were installed to provide additional roughing and cleaning capacity. The expanded plant has been operating since June 1993.

Jan 1, 1994

By Stuart M. Jones

This paper is derived from a document that compiles autogenous and semi-autogenous mill installations worldwide, through the end of 2010. Updates have been solicited from all major mill manufacturers and most have responded. A review of AG/SAG mills sold to date and in the last five years indicates that the AG/SAG mill market is very healthy and the trend is larger diameter mills with high diameter-to-length ratios. Ring motors (gearless) are the preferred drives for large mills even though the power transmitted through a gear drive is expected to near 11,000 kW per pinion in the near future.

Aug 1, 2010

By Yongqiang Liu

Background For refractory gold ores with high carbonate contents ! Acid pressure oxidation ! High acid consumption (if low sulfur content) ! Low oxidation efficiency due to CO2 in autoclave ! Problems with slimy precipitates (gypsum, etc.) ! Alkaline pressure oxidation ! Difficulty in fixing arsenic ! Low gold dissolution due to iron hydroxides/oxides coating ! High reagent costs Summary ! Carbonate and sulfide minerals had similar floatability across the pH range tested (5 - 12) using sodium oleate as a collector; ! Thioglycollic acid or citric acid depressed the flotation of pyrite but could not depress chalcopyrite when used alone; ! Combination of thioglycollic acid and citric acid had a much stronger depressive effect, and the mixture depressed chalcopyrite as well as pyrite between pH 8 and 11 without affecting the flotation of calcite and dolomite; ! Calcite could be separated from chalcopyrite and pyrite; dolomite could be separated from pyrite but not from chalcopyrite under the tested conditions.

Apr 1, 2005

By Manouchehr Oliazadeh

Talc is used as a gelating, diluting, antimould and softening agent in the pharmaceutical industries. This article reviews efforts made for producing pharmaceutical grade talc from a mine located in the west of Iran, called Masoud Abad. Reaching to the pharmaceutical standards, was the principal goal of the talc beneficiation. Methods such as flotation and leaching were applied to eliminate the impurities. Considering to the negligible Fe content of the ore, the major impurity was calcium carbonate and to some extent dolomite. Depressants used in froth flotation were dextrine, quebraco and sodium silicate. Variation of collectors such as oil petroleum, paraffin, stearic acid were also used. After optimization of the effective parameters, followed by two stages of flotation (rougher and cleaner) acid leaching process was also applied as a complementary step to remove the impurities. CaO impurities in the sample was 11.63% which reduced to 1.03% after two stages of flotation, and finally reached to 0.03% CaO by leaching process. This figure is more lower than the world pharmaceutical standards (i.e. <0.6 % CaO). The highest recovery was 55% at the best conditions. Other impurities (i.e.As, lead, heavy metals ,?) which were at standard levels, confirmed by the authorized health organizations.

May 1, 2001

By David F. Machuca-Mory

Modelling the geometry of a vein is a crucial step in resources estimation. The resulting models are used as mineralization domain boundaries and have a direct impact on the tonnage of estimated resources. Deterministic models are often built using time consuming wireframing techniques usually based on hand interpretation of the drillhole intercepts. Another approach consists in coding the drillhole samples by a function of their distance to the veins contacts. The coding is subsequently used for modelling the vein contacts away from drillholes. This is a more efficient approach and is able to provide a measure of tonnage uncertainty. The use of location-dependent variograms improves the modelling by incorporating local changes in the anisotropy of the vein structure. This combined approach results in more realistic vein models, particularly when the geometry of the vein has been are altered by folding, shearing and other structural processes. The approach is illustrated on a realistic case study.

Oct 1, 2009

By Chin J. Chang

Powder metallurgy (P/M) technique is known for its capability of producing near-net-shape parts. Its specific characteristics have resulted in low cost and eliminated of many secondary machining. However, more and more P/M parts do require additional operations to fulfil their complex geometric features and surface roughness. Many of the machining factors that influence the machinability of cast and wrought steel parts, such as cutting speed, feedrate, coolant, tool geometry and shape, are also considered in the machining of P/M components. The special feature in P/M structure is the porosity that can decrease the machinability and shorten the tool life. Different variables have been considered in the machinability of P/M materials. Densities and the free machining additive manganese sulphide (MnS), are the two main factors of material composition, which dominate the P/M machining performance. Machining process is the main method that generates the desired working surface of machining components. This paper will discuss the metallurgical structure of the powder metallurgy components, their physical properties and possible issues when encountering machining operation. Turning operation has been tested in this investigation to show the relationship between the cutting speed, federate, cutting forces and chip formation. This investigation also discusses how the free machining additive, manganese sulphide, affects the P/M structure and machinability in order to understand how the metallurgical structure can affect the machining operation and, how to obtain the optimised surface results.

May 1, 2001

By A. J. A. Janse

The Argyle diamond mine is located in the Kimberley region in the northern part of Western Australia at approximately 16°43&apos;S, 128°23&apos;E in subtropical bush/shrub savannah. The 46 ha (113 acres) pipe is situated in the Halls Creek mobile zone, which was cratonized about 1800 Ma ago and adjoins the Kimberley Block, a plateau upheld by Early to Middle Proterozoic rocks underlain by presumably Archean basement. The crater and upper pipe zones of the diatreme have been preserved within tilted and faulted strata of Middle to Late Proterozoic age. Its age of intrusion is about 1178 Ma. The diatreme rock is not kimberlite but olivine lamproite which was not recognized as a primary diamond host rock until 1978. The Kalumburu Joint Venture, a syndicate of local and overseas mining companies, started in 1972 to prospect the Kimberley Block with a modest annual prospecting budget of $125,000. Annual expenditure doubled each year and reached $1 million in 1975 after which CRA took over management of the newly-formed Ashton Joint Venture with an annual expenditure of $2 million (all figures are in current US dollars). The Argyle diatreme was found near the end of the eighth field season and after a total expenditure of $12 million. Annual production is about 34 million carats which places Australia as the world&apos;s number one diamond-producing country, although, because of the low value per carat, it is number four in terms of total annual value. The global cost required for a major diamond mine is between $300 and $400 million, with a discovery every 10 to 12 years. Because Canada contains large areas of prospective platform rocks underlain by Archean basement it is likely that a major diamond deposit does exist there. It is probable that it will be found before the end of this decade and exploration expenditure may be in the order of $100 million.

Jan 1, 1992