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By T. S. Heah, S. T. Bishop

"The Lorraine property is located in the northern Intermontane Belt, north-central British Columbia. The property is underlain by intrusive rocks of the Duckling Creek Syenite Complex, an Early Jurassic, alkaline phase of the Late Triassic-Early Cretaceous Hogem Batholith. Two significant zones of copper-gold mineralization have been explored to date; the Main zone (historical Upper and Lower deposits) and the recently discovered Extension zone. A combined geological resource was estimated in 1975 for the Main zone deposits at 10 million tonnes averaging 0.7% Cu, and from 0.1 g/t Au to 0.34 g/t Au. The Extension zone, still at an exploration stage, is estimated to be considerably smaller than the Main zone.Copper-gold mineralization at Lorraine Mountain exhibits characteristics suggestive of both magmatic and hydrothermal origins, and may be related to orthomagmatic-hydrothermal fluid flow contemporaneous with magmatism and development of migmatitic fabrics. A depth greater than the depth of formation of classical porphyries may explain many of the variations displayed by mineralization at Lorraine. These characteristics distinguish Lorraine from other members of the alkalic suite of porphyry Cu-Au deposits in the Canadian Cordillera.IntroductionThe Lorraine prospect is located in the Omineca Mountains, at the headwaters of Duckling Creek, 56 km northwest of Germansen Landing (Fig. 1). The property is centred at latitude 55°55'N, longitude l25°26'W on NTS mapsheet 93N/14W. The area is typified by mountains of moderate relief, with elevations ranging from 1150 m in the valleys to peaks of 2000 m. Valleys are U-shaped and blanketed by glacial till which give way to steep, talus covered slopes and sharp ridges. Outcrop is generally limited to elevations above 1600 m (Fig. 2)."

Jan 1, 1995

By S. Masambi, S. M. Bradshaw, C. Dorfling

The presence of iron in base metal leach solutions complicates nickel and copper recovery processes. Leach solutions are commonly purified by precipitating iron as hematite, jarosite or goethite. Fast oxidation of ferrous to ferric is an important step to remove iron by precipitation at low pH values. This study reviews the ferrous oxidation mechanism and reaction rates in chloride media using molecular oxygen. This work contributes to the development of a better understanding of iron precipitation processing steps, which continues to be an important matter in hydrometallurgy. The specific objectives were to investigate (1) the effect of key process variables (acid concentration, copper concentration, and temperature) on the rate and extent of ferrous oxidation by dissolved oxygen and (2) the oxidation mechanism by varying chloride and hydrogen ion concentrations separately. Batch oxidation tests were performed using synthetic chloride solutions containing 45 g/L Fe and 3 g/L Ni at 50 and 80°C, dissolved Cu concentrations (0.3 to 3 g/L) and HCl concentrations (2 to 4.6 M). Regression analysis indicated that the significance of the effect of different variables on the extent of oxidation achieved after five hours decreased in the following order: HCl concentration > Cu concentration > temperature. Increasing the HCl concentration from 4 M to 4.6 M caused the degree of oxidation achieved after two hours to increase from 58.3% to 98.1%. The catalytic effect of dissolved copper on ferrous oxidation was more significant at low acid concentrations. Experimental data suggest that rapid oxidation of ferrous is possible at high HCl concentrations for two main reasons. First, the formation of ferrous chloride complexes allows ferrous oxidation to proceed without the formation of superoxide as an intermediate product (the rate limiting step in dilute acid solutions). Second, the formation of ferric chloride complexes stabilises the oxidation product.

Jan 1, 2016

By Yve Fourmanoit

"This paper discusses the exploitation of shallow underground deposits in a very altered geological environment. The methods, their adaptation to local conditions, their evolution, dilution control and recovery of mineral reserves will all be dealt with.The recovery of the B zone crown pillar under 40 m of till that was originally saturated with water, is of particular interest.Future underground development and production start-up of the A-2 zone, organization of mining operations, mechanical services, group incentives and safety practices will also be given.IntroductionThe underground deposit of B zone is not very deep. A 300 m shaft provides access to the mine. In addition , exploitation sublevels are developed every 20 m and are linked by a service ramp. The overburden located on top of the deposit is 40 m thick and is composed of a very dense till.A detailed work plan was initiated to permit mining under the overburden. To do so, it was imperative that efficient methods be developed. At a time when an accident at the Belmoral Mine shook the mining world , it was a task to convince provincial authorities that it was safe.While B zone was being mined , development began in the A-2 zone deposit which has recently been put into production (Figs. 1 and 2).The intention of this paper is to provide an outline of the mining method s and their particular aspects. To do so, some of the mine services are described and a brief outline of the under ground department is provided."

Jan 1, 1990

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 P. W. Jensen

Gold deposits in Interior Alaska are commonly coincident in time and space with reduced, mid-Cretaceous granitoids. The True North deposit of the Fairbanks district is several km from such granitoids, but yields mineralization ages indistinguishable from the Fort Knox granite-hosted deposit. True North, however, resembles sediment-hosted epithermal Au, demonstrating the variability of intrusion-related Au deposits in metamorphic rocks. The True North deposit (14M tons at 1.5 gpt) is hosted within a carbonaceous thrust complex separating eclogite facies from amphibolite facies rocks, located > 5 km from mineralized granite outcrops. The ores contain anomalous Sb, As, and Hg (stibnite is the most common sulfide) with ankerite-sericite alteration in steep NE-striking ?feeder? shears and low angle thrust shears. Nearest hornfelsed rocks are located > 1 km away, and across several high-angle faults. Fluid inclusion data indicates trapping temperatures of ~250C from low-density, low-salinity fluids at P < 0.5 kb. Au was originally present as solid solution in sulfides; minable ores are restricted to the oxidized upper 100 m of the deposit. The considerable differences between this deposit and the temporally equivalent Fort Knox intrusion-hosted deposit indicate no one description adequately characterizes this deposit type.

May 1, 2003

By R. B. Rao, B. Das, J. K. Mohanty, D. S. Rao

Physical, mineralogical and geochemical characterization, along with some beneficiation approach to pre-concentrate the Ni content of the chromite overburden material of the Sukinda Valley, India, have been carried out in detail. Mineralogically, the overburden consists of geothite, quartz, hematite, magnetite and secondary silicates in varying proportions. A specific Ni mineral phase is absent in this material with the Ni and Fe being mostly enriched in finer fractions. Distribution of Ni and Fe in various size fractions and increase of Ni with increase in Fe (goethite content) indicate that goethite is the major Ni-bearing mineral phase. Quantitative estimation of goethite from chemical analysis and weight loss data indicates that the goethite quantity varies from 40% to 90%. Electron microprobe analysis (EPMA) indicates that Nirich and Ni-poor goethite are present in the overburden material. A sequential dissolution study indicates that more than 50% of the Ni are present within the goethite lattice as a lattice substitution. The results of pre-concentration studies (size classification, gravity, flotation and magnetic separation) indicate that Ni is richer in finer fractions. Further, the enrichment of Ni content is limited to 1.1% with 50% recovery by classification followed by hydrocyclone separation of the below 53 m m fraction.

Jan 1, 2000

By G. W. Poling

"Construction, equipping and planned operation of a new laboratory for engineering-teaching and research in coal and mineral processing is described. This facility will provide approximately 1860 m2 (20,000 ft. 2 ) of laboratory space with nearly half devoted to a new venture in coal preparation. Two high-bay, overhead crane-accessed areas will accommodate major unit operations for each of the ""coal"" and ""mineral"" laboratories. COAL PREPARATION FACILITIES:The coal facility provides three main functional areas:(1) Sample preparation (drying, blending, crushing, screening) and sink-float washability testing.(2) Unit operations U1igh-bay, serviced by crane).(3) Analytical laboratories.(l) Sample Preparation and Sink-Float TestingBreaking or comminution of coal will be accomplished using a rolls crusher, a hammer mill, a jaw crusher, or a pulverizer. Capacity to prepare coal at rates of up to approximately 4 tph will be provided by some of these units. Large diameter (~0.5m) baths of heavy organic liquids will be available for sink-float washability tests on coal samples of several kg. Particle sizes of up to 25 mm can be tested in these baths. Smaller scale washability tests on finer samples will be accelerated by centrifuges."

Jan 1, 1980

By Chris Larsen, Ian Orford, Mike Cooper

"An autogenous grinding circuit was commissioned at the Brunswick concentrator in October 1998. The autogenous grinding circuit replaced existing surface crushing plants, fine ore storage, rod and primary ball mills. The 8.54m diameter x 4.27m long (EGL) AN1 mill is driven by a bidirectional variable speed 6.3MW ABB wound rotor motor with Slip Energy Recovery (SER) drive. The AG mill was subsequently converted to SAG operation in May 1999 for a number of reasons. The mill is operated in closed circuit with horizontally inclined 660mm diameter cyclones to produce a circuit product P80 of 85 microns. Extensive test work and JKSimMet design modeling resulted in a production circuit operating closely within design expectations.The circuit is highly instrumented and controlled via a hybrid Programmable Logic Controller (PLC-5) and Fisher PROVOX Distributed Control System (DCS) that is in communication with the property information highway. An expert control system (Gensym G2) is used for control of the SAG mill load.Engineering and construction were completed in 28 months from the time of management approval for the project. The Brunswick AG/SAG mill circuit was commissioned on schedule at the $M 23.5 forecast budget."

Jan 1, 2001

?General Facts:Barrick?sFirst Mine in Peru Production started in November 1998 ?The Facilities:Open Pit, Primary and Secondary Crusher Valley Fill Heap Leach PadMerrill Crowe PlantRefinery Environmental Facts: High Rainfall - Average Annual Precipitation of 1,200 mm (47 inches) Rainfall Events: Common 20mm daily events High Altitude Site - Elevations range from 3900 to 4250m. Steep topography - Terrain ranges from 3:1 slopes up to cliffs. Water Issues: Sediment Control - Open Pit PAG Encapsulation - Waste Dump Stormwater control - Leach Pad Erosion Control Techniques ? Offsite impacts mitigation Social Issues: 11 communities surrounding the property Town of Huaraz

May 1, 2003

By A. H. Ransom, M. E. Robb

The Salmita Gold Mine is located 240 air kilometres northea t of Yellowknife on the east shore of Matthews Lake. Salmita is one of several gold showings which are regionally associated with the steep easterly dipping contact zone between mafic, intermediate and fe lsic volcanics to the west and overlying finely laminated argillaceous sediments to the east. The volcanic assemblage forms a narrow 2 m to 7 m wide north to northwesterly trend ing Belt of Yellowknife Supergroup Archaen greenstone rocks. The contact zone exhibits exhalite character and the increased sulphide content provides a di stinct VLF-EM , IP/Resisti vi ty and/or magnetic signature. Gold occurs at or near this contact in three main settings: associated with si liceous vein-like structures; with finely disseminated sulphides within ericitic altered upper fe lsic tuffaceous unit ; or with quartz in narrow shear zones within the overlying sediment . The B-Vein is of the former type with free mi ll ing gold occurring within a highly sil iceous zone having an average widtn of 1.69 m. Alteration is intense within the immediately su rrounding host lithologies and includes the development of si lica, carbonate, sulphide, sericite and chlorite with minor graphite, epidote. tourmaline and scheelite. The mineable zone above the 4th level (-200 m) ha an average strike of 160 m and is divided into three south plunging ore shoots separated by the regional Salmita Shear Zone in the south and by a low-grade structural hiatus in the north. Production from the B-Vein commenced in August 1983 based on a diluted in situ ore reserve of 140,000 t grading 26.6 git Au . As of December 31, 1985, a total of 131 ,537 t grading 27.8 git Au have been mined at a ra.te of 160 t per day and a 96.4 percent recovery, to produce 3,522 ,000 g (113,227 ounces) of gold . Current (January I, 1986) reserves total 45,500 t grading 37 g/t Au down to the 10th level (-500 m) .

Jan 1, 1986

By Arnfinn F. Prugger, Fritz F. Prugger

"The Saskatchewan potash beds are overlain by a number of water-bearing formations. There are water zones in the footwall as well. These aquifers created difficulties during the construction of shafts to the potash beds at 1()(}() m depth, and they pose a continuous threat to the safe and efficient operation of potash mines in the province.The first shafts to be sunk in Saskatchewan encountered major problems during shaft sinking. Of the 17 potash shafts started in the province, five had major water inflows during shaft sinking or after production startup, and one was abandoned because of water problems. The biggest obstacle encountered during shaft sinking was a poorly consolidated sand-shale-clay-siltstone unit commonly called the Blairmore Formation, which occurs at a depth of about 500 m. The initial problems with shaft sinking were overcome in the 19605by using European methods of freezing and the installation of cast-iron tubbings or a double-steel lining through water bearing horizons.From the start of production in the early 19605, operators have been aware of the ever-present hazard of potential water inflows into mine openings from permeable units in both the hanging wall (carbonates of the Dawson Bay Formation which occur 15 m to 50 m above the mining level) and the footwall (carbonates of the Winnipegosis Formation which occur 50 m to 150m below). This awareness has been reflected in conservative over-all extraction rates and mining layouts specific to local conditions. In spite of this consciousness, a number of water problems have occurred in operating mines. These have resulted in everything from simple increases in expenditure to the shut-down of one mine in 1987.In this paper a history of water problems in the Saskatchewan potash industry is given, followed by a summary of measures that are now commonly used to avoid these problems. A description of each inflow event is given, and the attempt is made to analyze the causes of each occurrence. Based on what was learned f rom each event, suggestions are made to prevent future water inflows into potash mines."

Jan 1, 1991

By A. M. Cann, C. I. Godwin

"The Kemess porphyry copper-molybdenum deposit occurs in the Intermontane Belt of north-central British Columbia. The deposit is in andesitic to basaltic flows and tuffs which may be correlated with the Savage Mountain Formation of the Upper Triassic Takla Group. Most known chalcopyrite and molybdenite is associated closely with a 500-m-diameter zone of intense quartz-sericite-pyrite alteration. A zone of supergene enrichment, up to 30 m thick, underlies a leached capping about 20 m thick. The age of mineralization and the surrounding granodiorite and quartz monzonite pluton is Early Jurassic. K-Ar and Rb-Sr analyses indicate that the mineralization and alteration is slightly younger (l82±6 Ma) than the granitic intrusion (KAr: about 205±6 Ma; Rb-Sr: 190±4 Ma). The low 87Sr/J6Sr initial ratio for the granodiorite is typical of plutons in the Intermontane Belt.IntroductionKemess is a medium to small porphyry copper-molybdenum deposit with significant gold and silver values (Pilcher and McDougall, 1976). This paper describes the lithologies and alteration at this property; geochronology by Rb-Sr and K-Ar methods defines the ages of mineralization and surrounding plutons to be Early Jurassic.The property, located in the Thutade Lake area of northcentral British Columbia (NTS Map 94E; lat. 57°04&apos;N; long. 126°44&apos;W), is about 425 km northwest of Prince George (Fig. 1). Closest road access is 35 km southeast of Kemess. Final access must be by float plane or helicopter. The British Columbia Railway right of way passes 70 km to the southwest of theproperty.A prominent gossan zone at Kemess was staked in 1967 by Kennco Explorations (Western) Ltd. following detection of geochemical anomalies during stream sampling the previous year. During 1968, 1969and 1971, Kennco completed 232 m of X-ray diamond drilling, conducted geological mapping and carried out geochemical sampling of silts and soils (Stevenson, 1969; B.C. Ministry of Energy, Mines and Petroleum Resources, G.E .M., 1969 and 1971). Getty Mines Ltd. optioned Kemess in 1975and 1976. They remapped the area and diamond drilled 12 BQ and NQ holes, totalling 2065 m (B.C. Ministry of Energy, Mines and Petroleum Resources, Expl. B.C., 1975)."

Jan 1, 1980

By M. A. Perras, M. S. Diederichs

"The Niagara Tunnel Project (NTP) is a 10.1 km long water-diversion tunnel in Niagara Falls, Ontario, which was excavated by a 14.4 m diameter tunnel boring machine. The excavated rock types included limestone, sandstone, siltstone, shale and mudstone. Based on observations and measurements the overbreak was divided into four zones. Approximately half the tunnel length was excavated through the Queenston Formation, which locally is a shale to mudstone. Three of the four overbreak zones were within the Queenston. Zone 1 includes the formations above the Queenston Formation, Zones 2 and 4 are border regimes with little or no influence from St. Davids Buried Gorge and Zone 3 is the area influenced by the gorge. Zones 2 through 4 can be generally summarized as having overbreak as the result of stress induced spalling. Zone 2 marks the transition from slabbing at the Whirlpool contact in the stress shadow to spalling towards Zone 3. The behaviour of Zone 3 is interrupted from the typical high stress behaviour due to the presence of St. Davids Buried Gorge. As the tunnel passed from the gorge the high regional stresses induce spalling, which creates the typical notch shaped overbreak geometry of Zone 4.Three modelling approaches were used to back analyze the brittle failure process at the NTP: damage initiation and spalling limit, laminated anisotropy modelling, and ubiquitous joint approaches. Analyses were conducted for three tunnel chainages: 3+000, 3+250, and 3+500 m because the overbreak depth increased from 2 to 4 m. All approaches produced similar overbreak geometries to those measured. The laminated anisotropy modelling approach was able to produce overbreak depths and chord closures closest to those measured, using a joint normal to shear stiffness ratio between 1 and 2. The back analysis results were able to produce sets of stress and strength inputs for different sections of the tunnel at the transition from overbreak Zone 3 to 4. The back analysis procedure demonstrates the importance of including the anisotropic stiffness in the numerical modelling approach to correctly capture the overbreak geometry and deformations around underground excavations."

Jan 1, 2015

By B. R. McEwen, A. N. Samis, J. H. Gray

"The Quebrada Blanca Copper Mine is an open pit copper project located at 4200 m elevation in the Chilean Andes. The deposit is a world-class supergene enriched porphyry with a mining reserve of 89 Mt of 1.3% copper. The supergene zone will produce 6.3 Mt of ore per year for more than 15 years to produce 75 000 t of cathode copper per year.This paper will review the project geology the computer model ling of the deposit, the mine planning process and the property production scheduling. With copper production .fixed by the onsite electrowinning capacity, the planning and scheduling aspects to optimize the project NPV will be discussed. Some of the logistics, equipment. personnel and process difficulties of operating a mine at 42()() m elevation will also be described. IntroductionProject DescriptionThe Quebrada Blanca Copper Mine is an open pit copper project located in the northern part of the Chilean Andes. The project is based on the supergene enriched portion of a porphyry deposit which has a mining reserve of 89 million tonnes at 1.3070 copper. Below the enrichment blanket, the protore represents a further resource.All ore will be leached, the higher grade material will be crushed and stacked for thin layer heap leaching, and the lower grade and oxide material will be dump leached as run-of-mine material. Solvent extraction and electrowinning will be used to produce 75 000 tonnes of cathode copper per year. Power will be supplied by a 40 MW on-site diesel generating plant with the waste heat being used for process heating. The leaching operation will operate in excess of 15 years and may be followed by open pit mining and conventional ore concentration of the underlying primary sulphide mineralization.Because of the dry location water will be piped 37 km from a ground water source at the Salar de Michincha. Water requirements will be approximately 1.5 million m3/year. The operation will eventually employ approximately 550 people on a 7-day-on, 7-day-off rotation with two twelve-hour shifts per day with an on-site camp. Busing will be provided to the port city of Iquique."

Jan 1, 1994

By Marc Ruest

A new instrumented cable bolt has been developed by the National Institute for Occupational Safety and Health at the Spokane Research Laboratory in Spokane, WA. Although various instruments are available to measure load distribution and magnitude along a grouted cable, this instrument is innovative in that it uses strain gauges internal to the cable as load-measuring sensors. The instrument has been successfully tested in the field at FMC?s Granger Mine, Meikle Mine, and Getchell Mine. A ?split-pipe? laboratory pull test was performed on each of three cable bolts instrumented with the new device. Each test consisted of grouting a 1.83-m-long cable in two 0.91-m-long sections of schedule 80 pipe. A numerical analysis was then performed in which laboratory boundary conditions were simulated and model properties derived from textbook guidelines. The loads calculated by the model were then compared to the measured loads.

May 1, 2002

By J. B. Yianatos, F. A. Contreras

Mineral carrying rates across the pulp-froth interface in the range of 1.4-3.2 tph/m2 and bubble surface coverages between 7 and 22 %, have been determined in large rougher flotation cells operated under normal conditions. To evaluate these variables, measurements of bubble size distribution, bubble load and size distribution of particles collected by true flotation, were performed at different concentrators. A model for estimating the bubble surface coverage at the pulp-froth interface level as a function of bubble load, Sauter mean bubble diameter and particle size distribution of the mineral collected by true flotation, was developed. Bubble loadings in the range 24-70 kg/m3, bubble sizes between 1.5 and 3.2 mm and superficial gas rates between 1.2 and 2.0 emfs, were measured in copper rougher flotation cells under normal operating conditions. Estimated bubble surface area fluxes were in the range 31 to 62 cm2/s/cm2- Results of carrying rates and bubble surface coverage, at the pulp-froth interface, were significantly lower than the maximum carrying rate for the collection zone, indicating that in this case is difficult to achieve maximum carrying rate at the pulp-froth interface level. The main constraints regarding carrying capacity limitations are then related to the froth transport characteristics (i.e. cell design, froth residence time, froth stability, froth recovery). Advances in Mineral Processing Science and Technology Proceedings of the 48th Annual Conference of Metallurgists ofCIM Sudbury, Ontario, Canada Edited by

Jan 1, 2009

By W. B. Boggs

A paper describing the Noranda smelter was presented at the 1930 annual meeting of the Institute (1). A lacer article appeared in the Trans-actions of the A.I.M.&M.E. in 1933. Since that time there have been many changes and improvements in plant operation and it is the abject of this paper to give an account of developments up to the beginning of 1939. The Noranda smelter was built originally for a capacity of 30,000 tons a month. At present, the average tonnage being treated is 110,000 tons a month, the tonnage of custom ores alone amounting to almost the total rated capacity of the original plant. This increase has been accomplished by increasing the capacity of the reverberatory furnaces from 700 tons of solid charge per furnace day in 1928 to 1,500 tons per day at the present time. In 1935, it appeared chat the limit of furnace capacity had been reached at about 1,300 tons of charge per day because of the short life of the sprung silica arches. However, the adoption of the suspended magnesite roof, which had been developed at the Flin Flou smelter of the Hudson Bay Mining and Smelting Company, overcame this difficulty and it is believed that the limit of the capacity of the present reverberarory furnaces has not been reached. A brief account of the changes to equipment in the various departments will be given, followed by a description of the more important features of present smelter practice.

Jan 1, 1939

By D. T. A. Symons, M. M. Kapnistos

"Gravity profiles were established over five known Middle Silurian pinnacle reef s in Lambton County, southwestern Ontario. The profiles were approximately 3 km long, with stations every 60 m. Four fairly recent technological developments were used which now make the gravity method relatively efficient and inexpensive compared to the techniques available two decades ago. These are: (1) a La Coste-Romberg model G gravimeter to measure the gravitational attraction with a sensitivity of 0.01 mgal; (2) a Sokkisha model B-2 semiautomatic engineers level to measure the station elevations with an accuracy of 0.5 cm; (3) a Huntec FS-3 hammer reflection seismograph to determine the overburden thickness to an accuracy of ± 7%; and (4) a modern microcomputer to process the data sets and mathematically filter the gravity values. Thus, two men can measure and analyze approximately 1 km of profile per day and obtain the residual gravity profile to an accuracy of ± 0. 05 mgal. Positive gravity anomalies of 0.17 to 0.72 mgal were observed along all five profiles that approximately delineate the dimensions of the known pinnacle reefs, which have an average height of 120 m, diameter of 1500 m and depth to top of 575 m. The relative ease of the gravity survey method makes it a cost-effective technique for hydrocarbon exploration in southwestern Ontario. IntroductionPinnacle reefs are important sources of oil and gas. Although the pinnacle reefs in southwestern Ontario are small, they may contain commercial pools of oil and gas because of their proximity to refineries. They are also useful for subsurface storage of gas to meet the winter surge in gas demand.At present, geophysical exploration for new pinnacle reefs is done primarily by relatively expensive seismic methods, which are also limited in southwestern Ontario by prohibitions on blasting. Gravity methods of exploration were attempted more than two decades ago in this region. (Pohly, 1954); however, the level of technology at that time made the method marginal in sensitivity and quite expensive. Several subsequent developments- gravimeters with greater sensitivity and operating ease, ""self-levelling"" levels for elevation control, hammer seismic reflection units for determining overburden thickness and portable microcomputers for efficient data analysis-have significantly improved the precision and cost effectiveness of the gravity method. Five known pinnacle reefs were selected in Lambton County, southwestern Ontario (Fig. 1), to test the effectiveness of integrating these developments to locate and define additional reefs. Isolated reefs were chosen to avoid the problem of the gravity effect s of one reef interfering with those of a neighbouring reef. The surveys were facilitated by choosing reefs with a concession road passing directly over them ."

Jan 1, 1980

By J. Peterson, D. D. Tannant, J. Todd

"A review of blast designs and improved blasting practices at the EkatiTM diamond mine is presented along with the results of three blast-monitoring experiments. Blast monitoring was undertaken to investigate blast damage mechanisms in the mine’s well jointed rock mass. Rock mass damage caused by production, pre-shear and wall control blasts was measured. EkatiTM uses 270 mm holes for production blasting, and 165 mm holes loaded with a decoupled charge for pre-shearing. The production blasts are loaded with bulk emulsion/ANFO blends. The mine plan involves using 30 m double benches. A 30 m high pre-shear is drilled and blasted prior to the production blasting that is done with sequential 15 m benches. This paper summarizes the monitoring equipment and data gathered to date. IntroductionThe EkatiTM diamond mine is located about 300 km northeast of Yellowknife, Northwest Territories, Canada. The mine is a joint venture of BHP Diamonds Inc. (80%) and geologists Charles Fipke and Stewart Blusson (10% each). The EkatiTM mine is Canada’s first diamond mine and BHP is the operator. The mine is accessed by air and by a winter ice road.The first kimberlite pipe mined at EkatiTM is known as the Panda pipe. The Panda open pit mine will have a total depth of 315 m (Fig. 1), a width of about 600 m, and a 50° overall slope. Initial ore production is 9000 tonnes per day.This paper summarizes the pre-strip blasting and the evolution of the production and wall control blast designs from 1997 to 2000. The initial blast designs and the rationale behind subsequent changes that took place are presented along with the techniques currently in use.The paper also briefly presents an experiment done to measure the rock mass response to blasting. Three blasts were instrumented with geophones, gas pressure sensors and time domain reflectometry cables. A brief summary of the instrumentation and data gathered is included."

Jan 1, 2001

By J. M. D. Wilson, W. Petruk, C. Coté

"The behaviour of minerals in the Mount Wright spiral circuit was investigated to find methods of improving iron oxide mineral liberation and recovery. The feeds, concentrates and tailings from rougher, cleaner and recleaner spirals in one mill line were sampled before and after the sizing screen openings were reduced from 2 mm to 1.6 mm. Sized fractions of each sample were analyzed with the MP-SEM-IPS image analysis system at CANMET to determine mineral quantities and liberation characteristics of the iron oxide minerals. Mass balance techniques were used to calculate stream recoveries of iron, silica, total iron oxide minerals, hema¬tite, goethite, quartz, ferromagnesian silicates, and liberated and locked total iron oxide minerals for each sieve fraction.The reduction of the sizing screen openings increased the ap¬parent liberation of the total iron oxide minerals in the screen undersize from 89% to 94% and raised the over-all iron recovery from 84.4% to 85.3%. Apparent liberation of iron oxide minerals in the cleaner sand also increased from 70% to 85%, but the – 75 µm fraction re¬mained the only fraction containing an acceptable amount of liber¬ated iron oxide minerals in each case. Due to limitations of spiral and grinding efficiency it is recommended that the cleaner sand be screened at 300 m, the +300 m fraction be reground, and the – 300 p.m fraction be rerouted to the rougher spirals.The amount of fine liberated iron oxide minerals (– 106 m) lost in the tailings was 46 TPH in mill line 3. This value represents 41 wt % of the free fine iron oxide minerals, or 7 wt % of the total iron oxide minerals in the spiral feed. Further work is needed to investigate the possibility of recovering this material by other methods.Middling iron particles (010 50%) and presumably free gangue in the size range –300+ 75 m tend to be recovered by spirals by entrapment."

Jan 1, 1990