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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 Anand Raman, Steve Dixon

"Industry practice for defining acid consumption is based on column tests. Freeport-McMoRan (FCX) has adopted a practice of leach operation using a defined pH or gpl acid in the raffinate based on the column test program. The acid consumption can be easily calculated using the leach operating parameters of SX-EW efficiency, lift height, bulk density, irrigation rate, primary leach cycle time, acid cure dose and acid concentration in the raffinate. The column leach test results and a column to heap efficiency factor are used to define the copper extraction. The maximum acid consumption will occur when zero acid is contained in the raffinate. The term Net Extracted Value (NEV) is to define the maximum extracted value relative to acid consumed.This paper presents the operating logic and method used to calculate acid consumption based on column tests and operating parameters.INTRODUCTIONIndustry practice for defining acid consumption is based on column tests. This may result in lower economic value if the tests are not designed to account for operating variables. A test program is described to assist in defining lift height and the influence of gangue on solution chemistry. A simple spreadsheet is presented that will use the leach operating parameters of SXEW efficiency, lift height, bulk density, irrigation rate, primary leach cycle time, acid cure dose and acid concentration in the raffinate to calculate the acid consumption for a proposed operation.A review of current leach operations solution chemistry will help to define acid consumption. This is recommended for all current leach operations.DISCUSSIONThe current practice at Freeport-McMoRan (FCX) is the use of column test results to define operating parameters. The monitoring of operational solution chemistry and copper production addresses any changes needed in operational parameters. This practice maximizes the value extracted from each tonne of ore leached."

Jan 1, 2008

By Jesse L. Duke

Outline ?Understanding the context ?The Business, political, social and legal environment ?Recipes for Success (three Ps) ?Develop effective Partnerships/relationships ?Plan carefully ?Actively Manage the ?Process? Business Environment ?Nunavut?Impressive mineral potential ?Emerging diamond industry ?Dominated by Juniors ?NWT ?Maturing Diamond industry ?Emerging oil and gas industry ?Yukon ?Alaska Highway Pipeline ?Junior-dominated exploration and development ?Growing acknowledgement in Canada of the importance of northern resources ?Shift of thinking: north-south infrastructure (vs. East-west)

May 1, 2006

By Tim Martin, Dave Tahija

"In April, 2014, Hecla's Greens Creek mill, near Juneau, Alaska, switched from sulphuric acid to the use of dissolved carbon dioxide for lead flotation circuit pH control. The Greens Creek mine uses paste backfill, in which mill tailings are mixed with cement and used to fill mined areas. Some backfill inevitably becomes part of mill feed during subsequent mining cycles, resulting in high feed pH levels and increased dissolved calcium in the lead flotation circuit. Sulphuric acid is traditionally used to lower pH in such cases but it produces undesirable side effects, including poorer lead-zinc separation and unstable froth. The amount of sulphuric acid used at Greens Creek was limited due to these effects, resulting in higher than target pHs when feed backfill content was high. Dissolved carbon dioxide can be added without these limitations so pH levels have been under much better control since the switch. Additionally, lead-zinc separation and froth characteristics have significantly improved.IntroductionThe Greens Creek Mine, owned by Hecla Greens Creek Mining is an underground mine located on Admiralty Island near Juneau, Alaska, producing silver, gold, lead and zinc in the form of lead and zinc flotation concentrates and a gravity gold concentrate (McDonald, Morgen & Hancock, 2012). The lead and zinc concentrates are produced in conventional flotation circuits preceded by a graphite removal flotation step.The underground mine utilizes cemented paste backfill, of which a certain amount is inevitably re-mined in later mining cycles and returned to the mill, mixed with fresh ore. The Portland cement used in the paste backfill is highly basic so that, from the point of view of the mill, the net effect is to introduce a good deal of lime to run-of-mine mill feed, producing a very high natural pH. This high pH needs to be lowered to obtain good metallurgical results."

Jan 1, 2016

By L. K. Walker, G. A. Kosick, P. D. Young

"MillEX is an information management system designed by mineral processors for mineral processors. The system utilizes ""cutting edge"" software technology to gather and store all automatically and manually recorded data from the plant operation and laboratories. This data is delivered to the desk of superintendents, metallurgists, technicians and managers, and permits the user to query, correlate, graph, report and present this data in minutes. In the past this task would have taken days and even weeks of tedious work to complete.MillEX allows the engineer/analyst to quickly develop a very thorough understanding of the complex relationships of process variables and equipment that exist in the plant, and their effect on product quality and quantity. The time and resources that are currently wasted on duplicating data entry will be an effort of the past.This paper describes the structure of MillEX and how it conforms to the needs of mineral processing personnel.INTRODUCTIONPicture a personal computer on your desk which will carry out the following commands:1. Plot a graph over the last six months, using weekly averages, of percent minus 200 mesh material in the primary cyclone overflow.2. Overlay this graph with a graph of concentrate grade versus recovery over the same time period.3. Retrieve a prior month end report, update all operating supplies tables and metallurgical performance with this months data, and embed the graphs created above.4. Edit text and print the current month end report. This is only one example of the type of power MillEX isbringing to the mineral processing industry."

Jan 1, 1992

By H. A. Schnell

"The Phase I Mill at Cluff Lake in northern Saskatchewan was designed to accommodate a very specific high grade uranium ore. As this high grade ore was depleted, construction was undertaken on a new mill circuit (Phase 2) to process a more conventional lower grade ore deposit.The Phase 2 Mill expanded the ore throughput to 800 tonnes per day. The new process employs crushing, grinding, leaching, counter—current decantation, solvent extraction and magnesia yellowcake precipitation and drying to produce a yellowcake concentrate.This paper discusses the process decisions required for this expansion and outlines the new mill circuit. Included is a review of the start up in August, 1984.1.,INTRODUCTION The Cluff Lake Uranium deposit is located 720 kilometers north of Saskatoon in northern Saskatchewan. A high grade deposit was discovered in 1970 with the first production taking place in 1980 after lengthy hearings. The original mill (or Phase 1 Mill) was specifically designed to accommodate this very high grade ore for a very limited plant life.Pilot plant work was carried out in 1981 on the other lower grade ores found in the immediate Cluff Lake area. Feasibility and process selection studies were then undertaken as the Phase 1 mill was nearing its end of life. These studies were completed in March 1983 when approval was given to construct the new Phase 2 mill."

Jan 1, 1985

By J. Kazakoff

"The success of gold recovery at the Jolu gold mill depends on many variables. The most important ones are the people, the ore body, plant design parameters, equipment selection and cost control. This paper will briefly hiqhlight the above, the main focus, however will be on the gravity gold circuit.The gravity circuit requires judicious operator attention in order to operate properly. In return the circuit gives immediate visual results for mill peopleIntroductionThe development of the Jolu property to the production stage was carried out quickly, efficiently and economically. The property became a gold producer ahead of schedule and under budget.The Jolu property is a joint venture between International Mahogany Corp., a 70% partner, and Corona Corporation, a 30% partner. Corona Corporation operates the Jolu Gold Mine, located 140 kilometers north of LaRonge and 520 kilometers north of Saskatoon."

Jan 1, 1990

By Mike Samuels, John Starkey

"This paper examines test results and grinding mill design for the recent Fortune Minerals’ NICO Project, located in the Northwest Territories. SAG milling and rod milling options are compared. Rod milling is clearly shown to be the best option because of the ore’s extreme resistance to breakage in a SAG mill environment.Surprisingly, the Bond Ball Mill Wi was 13 kWh/t for this ore. The Bond Rod Mill Wi was 20 kWh/t, whereas the SAG pinion energy to grind the ore to T80 1.7 mm was 30 kWh/t. The existence of ore with this ratio of SAG hardness to Bond BM Wi has not been previously observed. If not considered, this critical relationship may adversely impact the accuracy of mill design. This paper substantiates the significance of these findings to achieve successful grinding design on ore that is macro-competent during primary size reduction, but relatively normal during secondary grinding.INTRODUCTIONThe NICO Cobalt-Gold-Bismuth-Copper deposit is located 160 km northwest of the City of Yellowknife, Northwest Territories. It is the largest IOCG-type deposit (aka Olympic Dam-type) currently recognized in Canada with Mineral Reserves of 21.8 million tonnes contained within a larger 57 million tonne, lower-grade Mineral Resource. The project, owned 100% by Fortune Minerals Limited, has been assessed in a positive definitive feasibility study, subject to underground test mining and pilot plant testing, and is now in the permitting stage to develop a combined open pit and underground mine and process plant. An updated process flow sheet incorporates conventional crushing, grinding and bulk flotation, followed by filtration for shipment of a concentrate to a purpose-built refinery proposed to be built near Saskatoon, Saskatchewan. At the refinery, separate bismuth and cobalt concentrates will be produced following regrinding. The hydrometallurgical process methods include: pressure acid leach, precipitation, ion exchange, and electro-winning for recovery of cobalt as 99.8% cathode; solvent extraction and electro-winning of by-product copper cathode; ferric chloride leach and electro-winning for recovery of bismuth as 99.5% cathode; and cyanidation of the leach residues for recovery of gold as doré."

Jan 1, 2010

By Fernando Romero Lage

The purpose of this paper is to analyze detailed crushing and screening configurations and the interrelationship as they relate to mass flow balance, operating efficiency, plant availability and final production. The entropy of ore bodies limits the standardization of crushing and screening plants, however, consideration for a continuous and sensible assessment of the mass flow balance of the circuit will yield the best achievable results. Since energy costs have continuously risen in the last few years, it is imperative that these circuits offer low power requirements and low operating expenses. The paper will detail how the efficient implementation of continuous mass flow balance in the application of crushing and screening plants will generate quality improvement and process improvement, which will translate into bigger production and energy savings.

May 1, 2010

By C. D. Shackelford, J. F. Muñoz, M. Vauclin, P. Rengifo, A. Suazo

The results of a field-scale study involving in situ leaching of copper ore tailings using sulphuric acid as the leaching agent, injection wells to introduce the leaching solution and extraction wells to recover the copper-enriched pregnant leaching solution are presented. The study was performed at an abandoned tailings disposal site using a battery of injection and extraction wells. The medium was characterized through the determinations of physical, chemical, mineralogical, and hydro-dispersive properties based on laboratory and field tests, including a tritium tracer test, a pumping test, particle size analyses, and hydraulic conductivity tests. Numerous horizontal strata with different hydraulic conductivities were identified. A previously developed model was used to simulate the leaching of copper ore tailings. Although the study confirms the feasibility of using in situ leaching for recovery of residual copper from tailings deposits, the efficiency of the technology depends significantly on the medium stratigraphy and hydraulic conductivity.

Jan 1, 2001

By W. D. Gould, A. Kapoor, K. Morin, P. Bédard

This paper presents the results of a research study that is being conducted by the Mine Effluents Program, Mining and Mineral Science Laboratory (MMSL), Natural Resources Canada aimed at evaluating biotechnology processes for the treatment of nitrogen compounds such as thiocyanide (CNS) and ammonia (NH4-N) which are present in gold mill effluents. A sequencing batch reactor (SBR) technology, commonly used for the biological treatment of municipal and industrial effluents, was used in this study. In the SBR process, the micro-organisms were able to degrade CNS to NH4-N and NH4-N to nitrate (NO3-N) at operating conditions of two 12 h treatment cycles per day, with pH maintained in the 7.4 to 7.6 range, and at room temperature (approximately 21°C) and also at 12 °C. The end products of CNS and NH4-N biological oxidation were NO3-N and sulphate (SO4) that are relatively non-toxic. Partial removal of NO3-N was achieved by biological denitrification reactions in the SBR process. The SBR process effluent was measured to be non-toxic to rainbow trout based on the 96 h acute toxicity test. The microbial consortium isolated from the SBR treating a simulated effluent was able to effectively oxidize CNS and NH4-N to NO3-N in water samples (under batch conditions) collected at three mine sites located in Quebec, the Northwest Territories, and Yukon.

Jan 1, 2004

By G. Deschênes

The Eleonore property (James Bay district of Northern Québec), owned by Goldcorp, hosts gold within stockworks of quartz-tourmaline-arsenopyrite veins and veinlets contained within microcline (potassic alteration) and brown tourmaline replacement zones. The preliminary investigation to recover gold from drill core samples showed that flotation followed ultrafine grinding and cyanidation is the preferred option. Ultrafine grinding is used to liberate gold finely disseminated in sulphides. The flotation concentrate used in the investigation had a P80 of 11 microns and contained 75.5 g/t Au, 5.0 g/t Ag, 10.3% S, 0.07% Cu and traces of antimony (0.03%). The mineralogical characterization showed the presence of 65% gangue minerals, 23% pyrrhotite (hexagonal), 9.6% arsenopyrite, 2.2% pyrite and 0.08% chalcopyrite with trace amounts of galena (0.05%) and sphalerite (0.03%). Gold was in the form of native gold and electrum. A straight cyanidation conducted at pH 11.0, DO of 3-5 ppm, 2000 ppm NaCN, 20 C 72 hours extracted 97.0% Au and consumed 31.5 kg/t NaCN. The equilibrium was reached within 24 hours. Increasing the dissolved oxygen to 16 ppm reduced the consumption of cyanide to 22.0 kg/t with 96.7% Au extraction. A 16-h pretreatment with 2.0 kg/t lead nitrate, followed by a 24-h cyanidation reduced the cyanide consumption to 6.0 kg/t with a gold extraction of 97.5%. Increasing the lead nitrate addition to 6.0 kg/t reduced the cyanide consumption to 1.2 kg/t by passivating the surface of pyrrhotite. The gold extraction remained at 97.0%. It was possible to decrease the concentration of cyanide to 800 ppm NaCN without compromising gold extraction. The associated cyanide consumption was 1.2 kg/t NaCN. The passivation of gold grains was an issue when the cyanide concentration was decreased to 550 ppm NaCN. The gold extraction decreased to 92.2%. The average lime addition for all the tests was 7.4 kg/t and the average silver extraction was 80%. Efficient cyanidation of an ultrafine sulphide concentrate with only 800 ppm NaCN represents an advance in leaching of ultrafine sulphide concentrates.

Jan 1, 2013

By Andrea M. Ouchi

A major focus of ground control research presently being conducted by the Geomechanics Group in the University of British Columbia Mining Engineering Department is to incorporate weak rock masses (RMR76 less than 50) into existing empirical design relationships. An emphasis is being placed upon the updated span design curve/critical span graph originally developed at UBC. The original database has been augmented by mapping of one mine in Montana and of several mines within the Carlin and Getchell Trends of Nevada. In addition, a neural network analysis has been performed and modifications of the span design curves at the lower RMR76 range have been made. The common factor in all the above mentioned mines is a weak back and/or walls. In most cases, the ore zone is the weakest rock unit and must be dealt with to safely extract the mineral bearing rock. This work attempts to provide rock mechanic tools that will enable a mine operator to make economic decisions that will also ensure a safe working environment.

May 1, 2004

By Michael J. Kelly

IAMGOLD?s Westwood project is the site for the largest, by volume, bored vent raise in North America. When completed, it will be the fifth largest bored worldwide in a single pass. This paper will review the highest capacity raise drill system ever built (Redbore 100); as well as the technical difficulties in a single pass boring a 2,773 ft x 20 ft to 22 ft diameter raise. The difficult ground conditions and associated problems will be reviewed, as well as the solutions to overcome them. Use of the RVDS drilling tool to achieve .04% accuracy (14.5?) and the Geotube® water filtration system will also be covered.

Oct 1, 2010

By E. A. Lowe, J. W. Smith, C. W. Overton

"The operations of Bethlehem Copper are situated in the south-central interior of British Columbia at an elevation of 4900 feet above sea level. Main road and rail transportation systems pass through Ashcroft in the Thompson Valley, which is 28 miles by road from the minesite. The operation was the first low-grade open-pit porphyry copper mine in British Columbia.The minerals are chiefly chalcopytite and bornite, while the rock is principally feldspar and quartz. In the outer fringes of the orebodies some pyrite is present. Average grade of mill feed is 0.55% copper.History of Milling Operations Production commenced in December 1962 at a rate of 3,000 tpd in a plant that comprised two-stage crushing followed by rod and ball milling. Production ratings were increased to 6,000, 10,000 and 12,000 tpd by the addition of a further stage of crushing and more rod and ball mills. When the last expansion was completed tonnage throughput exceeded the 12,000 ton designed rate and was stabilized at 15,500 tpd.However, recovery did not meet expectations at the expanded rate because of coarse flotation feed. Capacity in the rod mill circuits exceeded the subsequent fine grinding capacity, resulting in poor liberation. Studies were initiated to improve recovery without decreasing throughput.Description of Operation Prior to Pebble Milling Crushing with primary, secondary and a closed tertiary stage proceeds at a scheduled 20.5 shifts per week. Fine ore screened on 1/2"" x 4"" apertures is stored in a 30,000-ton bin ahead of two 12.5' x 15' rod mills. Rod mill discharge joins secondary ball mill (one 10.5' x 14' and one 11' x 14') discharge for classification in a single 50"" cyclone which closes the secondary mills. Overflow from this cyclone is split between three third-stage ball mill-pump-cyclone circuits (two 11' x 14' and one 12.5' x 15'). Prior to addition of the two pebble mills the flotation feed was produced by this three-stage grinding system. Grinding criteria are shown in Table I."

Jan 1, 1975

By K. G. Stowe

"1. Introduction Mattagami Lake Mines and Orchan Mines operate zinc copper mines adjacent to one another near Matagami, Quebec, situated approximately 300 kilometres north-east of Rouyn-Noranda. (Note: Mattagami Lake and Orchan are in the process of being absorbed by Noranda Mines and will become Noranda Mines-Matagami Division). Both concentrators began operations in late 1963.The Mattagami Lake concentrator is capable of treating up to 3,500 metric tons per day of massive sulphide ore with a mineral composition shown in TABLE 1. The Orchan concentrator consists of two separate circuits - the Orchan circuit with a capacity of 1,050 metric tons per day and a custom circuit with a capacity of 825 metric tons per day. Throughout its operation, Orchan has custom milled ore from several smaller deposits in the Matagami district."

Jan 1, 1979

By P. W. Godbehere, B. W. Brooks

"This paper deals with the milling of slags from the Noranda Continuous Process, in the Horne Concentrator at Noranda, Quebec. The milling of copper slags, although novel in Canada, has been practiced in Japan, Finland and elsewhere for several years.In March 1973, the prototype Noranda Continuous Process furnace or ""reactor'' came on-stream, In this pyrometallurgical process (see Fig. 1), copper concentrates mixed with flux are smelted in a single furnace to form copper or a high grade copper matte, and slag. The conventional smelting processes of roasting, reverberatory smelting and converting have been incorporated in a single 17 ft. diameter by 70 ft. long steel cylindrical vessel (l, 2). The copper or matte is tapped from the bottom of the reactor into ladles and transferred to an anode furnace or to a converter. The slag is tapped into 40 ton capacity ladles, cooled, crushed, and conveyed to the concentrator. At the present time, the reactor is operated to produce matte. The matte slag* containing 4 - 6 per cent copper is milled using conventional ore dressing grinding and flotation equipment. A flotation concentrate assaying 35 - 40 per cent copper is recycled to the reactor. Crushing, conveying, grinding, flotation and tailing disposal equipment previously used for over 40 years for ore treatment has been modified to suit the requirements of slag milling. Slag milling is essentially the scavenger circuit of a pyrometallurgical process."

Jan 1, 1977

By Mario E. Rossi

The full economic evaluation of a mining project requires a block model that provides an estimate of the tons and grades of available resources. The model must be representative of the mineralization to be found at the time of mining for specified extraction periods and within acceptable error margins. In an operating mine, there may be a reasonable understanding and estimates of actual dilution and ore loss. This paper discusses the issue from the perspective of a project at a feasibility or preproduction level of development. Three types of dilution are discussed. Internal dilution (or change of support) is function of the spatial distribution of the mineralization and its relation to the operation?s selectivity. Geologic contact dilution is the unavoidable mixture at contacts or boundaries of material with different geologic characteristics, grades, or metallurgical performance. Operational planned and unplanned dilution should also be considered, sometimes incorporated by factoring the resource model, but more accurately by mimicking the grade control process. The resource estimation process always carries uncertainty, which is sometimes significant, and usually caused by several factors. These include quantity and quality of drill hole information; quantity and quality of geologic data; grade estimation methodology; dilution and mining recovery factors applied; etc. A key decision is to whether estimate as accurately as possible the in-situ geologic resource or to estimate the tons and grade of the material to be fed to the processing plant. The geologic modeling and grade estimation methods used are conditioned by this decision.

May 1, 2009

By Jan W. Merks

"The settlements between mines and smelters are based on the mass of metal contained in concentrate lots. The mass of contained metal is a function of the wet mass of a lot, and its moisture content and metal grade while the variance of the mass of contained metal is a function of the same variables and their variances. Uncertainties in the measurement procedure applied to estimate each variable are quantified in terms of precision and bias. The absence of bias and an acceptable degree of precision are the keys to effective risk analysis and loss control.ISO Technical Committee 183-Copper, Lead and Zinc Sulfide Concentrates, has developed standard methods for sampling and weighing, estimating precision and testing for bias, the determination of mass loss on drying, the calculation of the mass of contained metal and its variance, and various analytical methods for major and minor metals. The variance for wet mass is calculated from the precision of the scale while the total variances of moisture and grade are estimated from pairs of interpenetrating primary samples. The variance of contained metal can be calculated either from mine data or from smelter data but the risk associated with the settlement is largely determined by the absence or presence of bias and the variance of differences between exchange assays.ISO/TC 183 does not deal with precision and bias for exchange assays but is developing guidelines to determine realistic splitting limits. Given that the exchange of assays between mines and smelters accounts for a large part of the risk associated with the settlement, precision and bias for exchange assays should be measured and monitored. The need for umpire services can be reduced by applying statistical quality control (SQC).Calibrating static scales, moisture balances and analytical balances, bias testing of sampling methods and sample preparation procedures, estimating the variance of a single variable or the variance of a function that a set of variables defines, and monitoring precision and bias for exchange assays, are the quintessence of metrology, the science of measurement, when applied in mining and metallurgy."

Jan 1, 1995

By D. Komljenovic

This paper presents the use of queuing theory in truck fleet sizing in open pit mining operations. In the first part, a methodology is developed by which an optimal ratio between the size of shovels and trucks may be determined. The second part gives the fundamentals of the queuing theory and its application in analyzing shovel and trucks. The stochastic character of involved times (loading, truck cycle, etc.) is taken into consideration. Using the above theory, the optimal truck fleet size in actual conditions may be estimated. Minimal operating cost of the working system (shovels-trucks) represents the basic requirement of the analyzed process. The methodology was validated on Caterpillar front-shovel ? truck combinations.

Sep 1, 2004