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Everybody would agree that a good roadway leads to more efficient haulages by reducing haul times, smoother and safer operation for operator and less wear and tear on tires and equipment, thus improving availability and overall costs efficiency. There is a number of factors that come into play to achieve this especially when developing a mine. Excavations of Ramps and Headings: Operators should follow surveyors and layouts accurately while driving ramps. To maintain constant grades on ramps, which leads to better design control of water and easier grading of roadways. Dips and dives lead to needless over excavations and unnecessary cost and disposal of waste rock. When dressing a ramp some areas may require 2 to 3 feet of material to correct the dips and perfect to constant grade. You may ask yourself, is this all necessary? A small exercise is to follow a loaded haulage truck up ramp. At the first hump, grade change, notice the down shift and notice how far the truck travels to regain the up shift and his speed. The difference is remarkably noticeable when the grade is constant. All this leads to less driveline shock, less horsepower demand, better fuel efficiency and more productivity. Drill & Blast Patterns to control water for Ramp Headings: The drill pattern needs to allow for a ditch. A slight cross fall to form a ditch can easily be done with the lifters. The preferred side of the ditch is always the inside curve of the ramp as it has less tendency to erode roadbed material. On a level, the ditch is dictated by the sump location. A blasted ditch will eliminate water pounding and keep the roadbed dry. Water trapped in roadbed is the major cause of rollout and de-stability of roadbed. If the water is allowed to drain in the middle of the ramp it will cause erosion, movement and loss of fines and an increase in grading intervals.

Feb 1, 2008

By J. Jara

Oxygen, ozone and air were evaluated in the oxidation of ferrous iron in the presence of hydrochloric acid ranging from 0.1 to 4 M HC1. Stoichiometric coefficients were determined through mass balances of gases and ferrous titration. The molar ratio ferrous to ozone increases from two at 0.1 M HC1 concentration to six at 4 M HC1. For oxygen the same ratio is constant and equal to 2 and it is in agreement with other works. The stoichiometric coefficient of hydrogen ion was not defined due to the effect of hydrolysis of ferrous and in particular ferric ion on measuring acid concentration by titration with sodium hydroxide. Agitation, gas flow and pressure contribute in the utilization of gas. A technique to compare the gas efficiency at several conditions is discussed and it is based on the relationship between oxygen and ferric during time. There is a strong compromise between gas efficiency and ferrous conversion. Oxygen efficiency decreased with time due to ferrous depletion. The gas efficiency ranged from 3 % (low level of agitation) producing 25 g Fe(III)/h to 12 % at 1000 rpm with 60 g Fe(III)/h converted. At 405 kPa absolute the oxygen efficiency was in the range of 80 % producing 300 g Fe(III)/h. The replacement of chlorine with oxygen to enrich molybdenite concentrates by reducing copper to less than 0.2 % was evaluated. Oxygen can be employed in batch or continuous operation and it can treat molybdenite concentrates with high variation of copper content. The dissolution rate is mainly controlled by oxygen addition in which utilization of oxygen plays an important role. Improvement on safety and environmental conditions is feasible by using oxygen instead of chlorine gas. The economics indicate good opportunities to implement oxygen without significant modifications to the actual circuit and enough benefit when oxygen is well used.

Jan 1, 2007

By Angela L. Paveley

Economic viability of base metal mining is complicated by a volatile market that can cause fluctuations in the growth and stability of the mining industry. Nickel is mined primarily for use in manufacturing industrial stainless steel and plating materials. It is essential that nickel production be sufficient to respond to current and future market demand. A Post-Pillar cut and fill mining complex is currently producing nickel sulfide in Northern Manitoba. From 2006 this area has shown a steady increase in the annual production, rising from 11 million pounds to 25 million pounds in 2010, and accounting for 33% of the total production for the Manitoba Operations. It is expected that fine-tuning the mining process can safely enhance ore recovery, improve production consistency, and increase economic viability of the deposit. The Post-Pillar Optimization study was initiated by the local technical and operational departments at the mine and analyzed from July 2010 to March 2011. The purpose of this study was to determine if portions of non-recoverable pillars can be slashed without jeopardizing pillar strength. This ideal case would optimize ore recovery and reduce variability in the ore mining rate. Numerical modeling and geo-mechanical material properties analyses, completed by the rock mechanics group, define the maximum drift width at 32 feet and the minimum pillar width at 18 feet. Thus, it was suggested that up to 3 feet could be slashed from two sides of the standard 24 foot pillars without damaging ground integrity. The field data collected indicates that additional ore could be recovered without compromising the safety of the workplace and the integrity of the complex. Consequently, technical and operational groups work jointly to achieve pillar optimization. Approximately 500,000 pounds of nickel was extracted from the complex via pillar optimization from July 2010 to March 2011. Economic analyses show the pillar optimization method developed for the mining complex is a proven, value-added opportunity. The mining complex remains a safe workplace and has retained ground integrity.

Nov 1, 2011

Unplanned open stope hanging wall dilution is a significant cost for many open stope mining operations. Significant advances in empirical and analytical approaches for estimating stope dilution have been made, however, many important factors are still either ignored or assessed in purely subjective terms. There are many hard to quantify factors that simultaneously influence opening stability and dilution. This creates a significant difficulty for determining the influence of a single variable, such as exposure time, on opening stability. Open stope exposure time has been found to be one of the important factors that influences overall stope stability and dilution. A study is currently being conducted to quantify many of the factors influencing open stope dilution. The assessment of a large comprehensive case history database of Hudson Bay Mining and Smelting Co. Ltd. (HBMS) operations forms the basis of this study. This paper presents the findings of a study on the influence of stope exposure time on open stope dilution. The study is based on applying existing empirical stope dilution prediction methods to the extensive HBMS database. Minimising stope mucking time has been recognised as a method to significantly reduce stope dilution. Factors which influence hanging wall dilution, such as drill and blasting methods as well as undercutting the hanging wall, have been applied to reduce the scatter in the database. Empirical design techniques based on rock classification attempt to relate opening stability to ?stand-up time?, however, these approaches have not frequently been applied to the mining industry. Data from a ?stand-up time? graph have been interpreted in terms of an equivalent average depth of hanging wall slough and compared to the HBMS database findings. To add to the existing HBMS database, several stope case histories have been taken from Geco mine. These cases include data from very large open stopes with long exposure times. Data from all three sources suggest that hanging wall slough increases between .14 and 0.3 metres per month for a rock mass Q? value between about 4 and 13.

May 1, 2003

By A. Fyson, M. P. Smith, Margarete Kalin

"Drums containing 0.07 m3 of pyritic waste rock from a zinc/copper mine, with and without Natural Phosphate Rock (NPR) additions (40 kg/t waste rock) were set up outdoors. The effluent chemistry has been monitored over a period of 709 days. In the presence of NPR, the cumulative acidity discharged from 7 out of 8 drums was 80 % less than that discharged from drums without NPR. Mixing NPR throughout the drums was more effective in reducing acidity than placement of the NPR in a layer in the drum. NPR treatment was more effective for 'fresh' waste rock than for material which had weathered for more than 4 years. It is concluded that application of NPR to pyritic waste rock could greatly reduce acidity generated.IntroductionWaste rock dumps are significant sources of acid, generated by the oxidation of sulphides. Oxygen and water supply are oxidation rate-controlling factors, and oxygen supply is the reaction-limiting component. If secondary minerals can be formed on the surface of the rocks, pyrite oxidation and therefore acid generation may be reduced or inhibited. Neutralizing agents such as limestone do not prevent the oxidation of pyrite but treat the resultant acid mine drainage (AMD) or acid rock drainage (ARD).Meek et al. (1991) determined in pilot scale experiments that apatite mixed with tailings was more cost effective than liming or use of PVC liners to reduce oxygen and water penetration into the tailings. Phosphate is not acting as a neutralising agent but it creates a chemical condition on pyritic surfaces which prevents reactions of the oxidising agents. Precipitation of principally iron phosphates physically isolates the pyrite surfaces from further oxidation (Ziemkiewicz 1990; Evangelou 1994). The chemistry of the phosphate effects has been described by Spotts and Dollhopf (1990). Secondary mineral formation was evident in the leach pile at Gibraltar when leaching was impaired (Scott, 1991) indicating a link between oxidation and the mineral-formation process. This paper reports on effluent characteristics when waste rock is generating acid in the presence of NPR."

Jan 1, 1995

By T. S. Grewal, S. C. Gupta, H. Natarajan, T. W. McQuade, N. Nejibulloev-Mamedov

Cigar Lake is one of the world’s largest high-grade uranium mines. The ore mined at Cigar Lake is milled at Orano Canada Inc’s (Orano) McClean Lake mill. The average Cigar Lake ore uranium grade since the start of mining activities in 2014 is 17.8% U (range 8– 28%). To achieve budgeted production of 18 M lbs U3O8/yr the required average feed tonnage to the mill is about 45 kt/yr. Arsenic, occurring as arsenides and sulpharsenides, is one of the major elements of concern associated with Cigar Lake ore. The leaching process is a sulphuric acid leach under oxidizing conditions at near atmospheric pressure. The presence of arsenic in the ore consumes oxidant in leaching as well as ferric sulphate in tailings preparation for arsenic sequestration. The average arsenic grade since the start of mining at Cigar Lake is 0.57% As. Further, the ore is hosted in a matrix of sandstone and clay, which results in poor settling in CCD. The average uranium grade for the balance of Cigar Lake Phase 1 ore is expected to drop to 12.5% U and the average arsenic grade is expected to increase to 1.6 % As, with an annual average peak of 2.5%. Therefore the tonnage throughput required to sustainably meet the production target will increase to 55 kt per year. Orano partnered with Hatch Ltd. (Hatch) to perform a gap analysis study to identify key circuits and equipment in the mill that would require debottlenecking. A METSIM mass balance model of the McClean Lake mill was generated and validated with actual mill operating data. Various feed scenarios were then studied, and the outcome was a traffic light table, which highlighted feed scenarios that exceeded installed capacity of major equipment and circuits. The results provided the basis for the PDC for the pre-feasibility study. This paper will outline the results of the gap analysis, the mass balance model, the feed scenarios and the study undertaken to prepare the McClean Lake mill for the change in ore quality from Cigar Lake.

Jan 1, 2020

By J. Holder

"A key barrier to vertical mineshaft construction is the construction access required for workers. A suspended work stage is used to provide this access for the mechanical, electrical and structural applications and other tasks. Work stage design is complicated considering the numerous types of construction activities that are executed simultaneously, the confined environment in which the work stage operates, the significant corrosion attack that is present, the weight limitations imposed by the mechanical hoist systems, and the risks associated with worker safety. Also complicating the design, these structures and their construction environment can vary with time. Hence, they can be described as 4D structures. The author has discovered significant synergies and advantages with integrating engineering design with detailed three-dimensional (3D) fabrication design. The significant reduction of minor design errors needing field correction will be underscored. The tender process is also accelerated since the design output is fabrication shop drawings including a detailed quantity takeoff output which lists each component required and its specific quantity. The paper presented will document the numerous advantages to the project using this approach. INTRODUCTION A key barrier to vertical mineshaft construction lies in the access for the workers to execute the construction. Suspended work stages are used to provide this access for the mechanical, electrical and structural applications as well as other tasks. The design of such work stages is complicated considering the numerous types of construction functions that need to be executed simultaneously, the confined environment in which it operates , the significant corrosion attack that is present, the weight limitation imposed by the mechanical systems for suspension, and the risks associated with worker safety. Also complicating the design, these structures and their construction environments can vary with time. They can be described as 4D structures with the fourth dimension being time. This design is a time consuming process. The post tender award is meticulous and requires strong teamwork between the steel fabricator and the consulting engineer. This teamwork may or may not be present and is dependant on the experience and relationship the team members have working together as well as other factors."

Jan 1, 2019

By Gérald Riverin, Bernard Boily, Daniel Bernard

"The Donalda gold deposit is located in the southern part of the Noranda camp, approximately 2 /em east of the Horne deposit. Combined past production and current reserves total over 1.5 Mt grading 6.9 grams of gold per tonne.The deposit consists of two main veins which are both narrow but persistent structures. The No 1 Vein, hosting about 80% of the gold ore, consists of an EW-striking banded quartz vein dipping on average 25° to the south. Although variable in thickness, it averages lessthan1 metre. TheNo2 Vein, located 300 m below the No 1 Vein, is similar to it except that it dips 50° south. Gold occurs mostly as fine native gold disseminated in quartz, with 2-3% pyrite and trace amounts of various sulphides and tellurides. The wall rocks are only weakly altered and mineralized with pyrite, hematite and sericite.Structurally, the Donalda deposit is centered on the Horne Creek fault (HCF) where the HCF is intersected by the ENE-trending DonaldaandCentralfaults.NorthoftheHCF,theNo1Veinishosted by the Quemont Rhyolite, whereas south of the HCF the vein is hosted by magnetic basalts and diorites of the Wilco Andesite Formation. The No 1 Vein extends across the HCF with only minor displacement, whereas the No 2 Vein has not yet been traced south of the HCF.A preliminary structural analysis of the mine shows that the HCF, which strikes 275° and dips 70° north in the mine area, has a reverse movement a lid that the flat dipping No 1 Vein is a tension vein. Both probably developed in response to the same stress conditions. Although the absolute age of the Donalda veins is not known, structural relationships indicate that they formed during the latest movements of the HCF, but prior to the intrusion of the Proterozoic D' Eldona Diabase. The occurrence along the HCF of gold-rich synvolcanic massive sulphides and much younger vein-type gold deposits (such as Donalda) indicates that the HCF was a very long-lived, deep-seated (?)structure which has controlled the circulation of gold-rich fluids over a long period of time."

Jan 1, 1990

By D. R. Schmitt, Q. Jia, R. Kofman

"Tests were carried out on clear glass blocks containing borehole aligned with and at an angle to applied principal stress states in order to study the growth of drilling induced tensile fractures. The samples were illuminated such that photoelastic fringes could be video recorded during the measurements; this was useful in assessing states of stress and in determining fracture initiation. Axial drilling induced fractures were observed at the azimuths of the applied uniaxial stress as would be expected from the concentrated stresses. En echelon fractures were observed from the borehole deviated at 45° to the direction of the compression. These preliminary and currently qualitative results show that glass blocks serve as a useful test medium for studies of stress concentration and drilling induced fracture creation. INTRODUCTIONWith the increasing exploitation of unconventional oil/gas reservoirs using deviated drilling in recent years, the effects of the in-situ stress on the success of energy production have become more critical than ever before. Knowledge of the state of stress in the earth enables us to determine well trajectories, well locations, hydraulic fracturing design and production rates, but knowing the crustal stress is always challenging. Strictly, it can only be obtained to varying quantitative degrees by accessing the earth through drilling. Failure of the rock at the borehole wall produces borehole breakouts (e.g., Bell & Gough, 1979; Haimson & Herrick, 1986; Zoback et al., 1985) and drilling-induced tensile fractures (DITF) (e.g., Aadnoy, 1990; Aadnoy & Bell, 1998; Barton et al., 1997; Brudy & Zoback, 1999; Davatzes & Hickman, 2010; Peska & Zoback, 1995). Thus, it is important to understand the failure mechanism of the rock in the vicinity of the borehole. A number of closed-form expressions have been developed that allow the stress concentrations to be calculated in the vicinity of a borehole including Kirsch’s (1898) 2-D plane strain solution and 3D formulae for elastically isotropic (Hiramatsu and Oka, 1962) and anisotropic (Lekhnitskii, 1981; Amadei, 1983) continua. These solutions have been applied in numerous studies (e.g., Ong, 1994; Ong & Roegiers, 1996; Vahid & Ghassemi, 2011; Jia & Schmitt, 2014; Ito et al., 2001) to analyze the failure mechanisms of drilling-induced tensile fractures (DITF) on the borehole wall."

Jan 1, 2015

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 C. Mattich, Å. Holmström, L. Hedström

"The fluidized bed at Division Ministro Hales (DMH) division of Codelco in Calama, Chile is currently the largest operating roaster for partial roasting of arsenic rich copper concentrates. The roaster converts the Mina Ministro Hales (MMH) enargite rich (Cu3AsS4) copper concentrates into calcine with a low arsenic content that is suitable for mixing with other copper concentrates feeding a flash furnace. The plant sets a new standard for the processing of arsenic rich concentrates, both from the quality of the calcine as well as the environmental performances. The process technology and expertise are based on the solid practical experience of treating arsenic rich concentrates from both the Boliden smelter and Outotec.The partial roasting process is located next to the mine and concentrator and includes a sulphuric acid plant with auxiliary equipment. The produced calcine has an arsenic content significantly below target and the residual sulphur meets the requirement of the flash furnace. The plant operates with high efficiency cyclones resulting in a low amount of calcine passing to the hot electrostatic precipitator (HESP). The small flow of HESP calcine dust can be combined with the low arsenic calcine and still meet the flash furnace feed arsenic requirements.SO2 formed during roasting is converted into sulphuric acid, while volatized arsenic is recovered in the wet gas cleaning section of the acid plant along with a large proportion of antimony present in the MMH concentrate. Arsenic is stabilized as calcium arsenate and sent to disposal at a specially prepared site.The present production rate (September/October 2015) is close to or higher than the original maximum design capacity of the plant.INTRODUCTIONMineralogyThe mineralogy of an ore and concentrate is important, since there is a direct link between roasting performance and concentrate mineralogy. The MMH concentrate has a good combination of minerals that makes the roasting process efficient and flexible. The main minerals are pyrite, enargite and chalcocite (Figure 1). The inert content is about 15% and consists mainly of silicate and alumina based minerals. The lead content is very low and this is desirable, since the presence of even a small amount of lead will cause a substantial increase of sintering during roasting."

Jan 1, 2016

By J. W. Smith, J. Van Dusen, A. Papachristodoulou

"As part of a study to evaluate factors affecting tankroom air quality, the mechanism of aerosol generation from bursting bubbles in electrowinning operations has been investigated. Typically 2 to 3 mg/m2.s are produced in zinc tankrooms. Analysis shows that the rate of production of film drops (geometric mean diameter - 2 urn] increases with decreasing bubble size, and that jet droplets (geometric mean diameter - 20 um) are not produced with gas bubble larger than - 4 mm. Both mechanisms of drop formation and current efficiency are affected by surface active agents, with an over-all reduction in emission of up to 60% to 70% being achievable. A source control system which does not adversely affect electrolyte properties is always preferred, and coalescence of all gas bubbles in to a continuous stream of gas at a single point on the electrolyte surface would eliminate aerosol emissions. This objective was the basis for design of a coalescence device which, in laboratory tests, reduced the emission rate by over 95%.The control device fit s over the electrodes in the electrowinning cell and is partially immersed in the electrolyte. A number of changes in the design which resulted in significant improvement in the performance of the device are reported.Field tests have shown that only about 75% of the aerosol below 10 um can be eliminated by coalescence. The remainder (up to 95% of the total) is controlled by coalescence of droplets, by impaction on the exposed collector surface and gravitational settling.The device does not interfere with operation of the tankroom, handles the effect of surges in power during harvesting, and is simple and inexpensive to maintain. IntroductionDuring the electrowinning of metals (such as zinc and copper) gases are evolved along the electrode surface forming small bubbles. The bubbles travel upwards and burst on the electrolyte free surface generating, in this way, small droplets. Because the electrolyte is often highly acidic or contains toxic materials, the whole process can result in the formation of an aerosol in the atmosphere above the electrowinning cells which poses a serious health hazard to workers in the area, and corrodes equipment."

Jan 1, 1988

By Libao Zhu, David J. Williams, Nazife Erarslan

"Statistics show that the drawbell in underground mines is usually subjected to significant damage by impact loading due to falling ore. Drawbell damage causes a large quantity of residual ore to remain on the drawbell, thus leading to a huge waste of ore resources. In this study, the Cavity Monitoring System (CMS) and Surpac software were employed to visualize the 3-D damage. Also, the Discrete Element Method (DEM) was employed to perform simulations of the falling ore to investigate the damage evolution of drawbells under repeated ore impact and obtain the dynamic damage accumulation mechanism. The reason of drawbell damage was found due to the micro-shear cracks. The numerical simulations of a drawbell damage process under repeated impact loading showed the main failure type was shear failure and this type damage was found increasing with more ore falling. INTRODUCTIONDrawbells are an important part of underground mines and play a significant role in ore recovery in terms of safety, economy and efficiency (Brunton et al., 2010; Hustrulid & Bullock, 2001; Trueman et al., 2002). In most cases, blasted ore is gravitated along the drawbell and finally arrives at the footwall of the drawpoint. However, some of the blasted oversize ore blocks clog the drawbell road and those ore pieces are called hang ups (Penswick, 1997). However, any damage on a drawbell is believed to be the reason for these hang-ups. In the literature, researchers have reported that the stability of the draw-point is influenced by the stress distribution above and near the drawbell (Butcher, 1999; Neverov & Neverov, 2013). Much research in the literature has been done on draw-points in underground mines. In fact, rockfalls cause very serious damage on a small impact area in a very short time (Volkwein et al., 2011). Hoek (2000) pointed out that the blasted ore causes progressive ravelling and even collapse of the drawpoint, thus the ore in the stope is abandoned. In this paper, the Cavity Monitoring System (CMS) was employed to obtain the damage profile of a drawbell due to the impact of rockfalls and to confirm the residual blocking ore on the damaged drawbell. Then, the drawbell damage evolution due to repeated falling ore was simulated by the Discrete Element Method (DEM)."

Jan 1, 2015

By G. W. Toop

"The Cominco smeller in Trail, British Columbia, is a large integrated lead-zinc production facility. In the plant description, emphasis is given to the pyrometallurgical operations including zinc concentrate roasting, lead singer-blast furnace processing and silver refining. Production figures for the main metal products - zinc, lead, silver and gold - are provided.IntroductionIncorporated in 1906, Cominco Ltd. is an integrated natural resource company with principal activities in mining, smelting and refining, mineral exploration, fertilizer production and electrical power generation.Cominco is one of the world&apos;s largest producers of zinc and lead, accoW1ting for about 10070 of mine production in the western world. The company also produces and sells copper concentrate, copper sulphate, copper arsenate, sodium antirnonate, silver, gold, cadmium, bismuth, indium, and germanium dioxide and a range of fertilizers and sulphur products.Zinc OperationsRoastingThe roasters currently treat about 1400 t/d of zinc concentrates. Over half of the total concentrate treated originates from the Red Dog Mine in Alaska. During the summer, concentrate is shipped from Alaska to Vancouver where it is loaded into rail cars and hauled inland, transferred to trucks and delivered to Trail. The remainder of the roaster feed is made up of Sullivan concentrate and various custom concentrates. These arrive either by train or by truck.There are three roasters available for roasting the concentrate. The two major roasters are Lurgi fluid bed units. Each Lurgi roaster can process concentrate at about 1:7 tonnes per hour (tlb). A smaller Cominco suspension roaster with a capacity of 10 tlb is also used. The flowsheet for the high temperature section of the fluid bed circuit is shown in Figure 1.Concentrates fed to fluid bed roasting are conditioned with 7 1/min water to give 8% moisture in feed. Two slinger belts throw the mixture into the fluidized bed. The fluid bed roasters are 10.4 m in diameter with a bed depth of 1 m. Grate area is 84 m2• Normal operating conditions include bed temperatures of 930°C to 950°C, fluidizing air at 750 Nm3/min and a wind box pressure of 2000 mm H2O."

Jan 1, 1994

By Frank Ebbutt

Introduction The property of the Britannia Mining and Smelting Company, Limited, is situated on the east side of Howe sound, about six miles from its head. It is roughly thirty miles by water from Vancouver, one of the important ports of the Pacific and the industrial and financial centre of British Columbia. Howe sound is navigable for ocean-going vessels and Britannia Beach is a daily port of call for the commodious boats of the Union Steam-ship Company, sailing from Vancouver. Few mines are so ideally situated or have so many or varied natural advantages. Copper was first found on Britannia mountain in 1888, by Forbes, but it was not until 1898 that five claims were staked here by Oliver Furry, a trapper. It is from these claims that the major production during the past thirty years has come. In 1900 the Britannia Copper Syndicate was formed and development of the property was commenced. In 1903 the late Grant B. Schley obtained control, and through his efforts the Britannia Mining and Smelting Company was organized in 1908 as a subsidiary of the Howe Sound Company. Since that time, a persistent and progressive policy has been followed, which has resulted in a prolonged and profitable operation. The country in the immediate vicinity of the mine is marked by strong relief, with mountains rising abruptly from tide-water to elevations of 5,000 feet and more. Two broad transverse valleys deeply dissect the area. These are the valleys of Britannia creek and Furry creek. The region has under-gone prolonged and relatively recent glaciation. Over much of the surface there is deep overburden and dense vegetation. The geology of the mine and vicinity has been described in reports by S. J. Schofield (1), H. T. James (2), J. G. Barry, and others (3). The present writer is in accord with much of this literature on Britannia, but during recent years, extended observations made underground have shown that the &apos;quartz porphyry&apos;, or &apos;mine porphyry&apos;, of several of the earlier writers is not the host rock to the commercial mineralization; and he contends that the mineralization occurs in highly altered phases of a complex series of ruffs, flows, and pyroclastics and their autoclastic metamorphosed products.

Jan 1, 1935

By T. Robson

"A multi-seam underground mining operation employing mechanized room-and-pillar methods provides 35 per cent of Smoky River's annual ROM coal production (2 million tonnes) from the Smoky River coal field. Surface mines provide the balance. Tectonic activity, which has resulted in deformed coal seams and a rugged relief to the area, has provided natural boundaries to discrete mining areas. Coal transportation to the central preparation plant incorporates a 3 km long underground conveying system. The company is committed to efficient resource recovery and a safe and healthy working environment. Introduction The Smoky River Coal operations (SRC) are centered in the Smoky River coal field located in west central Alberta, some 360 km west of Edmonton and 20 km from the Town of Grande Cache (Fig. 1). Combined mining operations produce 2 million tonnes per year run-of-mine coal from which 1.65 million tonnes of high quality, low volatile, coking coal are obtained. The underground mining operation, employing mechanized room-and-pillar methods, provides 35 per cent of SRC's annual ROM coal production (2 million tonnes). Surface mines provide the balance. Tectonic activity has resulted in deformed coal seams and a rugged relief in the area. This has provided natural boundaries to discrete mining areas. Background Exploration of the Smoky River coal field took place over four decades commencing in 1926. McIntyre Mines Limited (parent to Smoky River Coal Ltd.}, commenced development of the property in 1968 having secured a long-term contract for the delivery of coal to Japan. Production of underground coal at Smoky River commenced toward the end of 1969 and has been continuous on the property since that time. Originally, two underground mines were opened. Original planning envisaged the use of room-and-pillar techniques to develop blocks of coal for subsequent extraction by longwall retreat mining. Following initial development, two longwall units were established, the first in November 1970 in No. 5 Mine and the second in February 1971 in No. 2 Mine. Difficulties were experienced on both units, and production commitments could not be achieved. The longwalls were discontinued, and mechanized room-and-pillar was subsequently used for all underground mining."

Jan 1, 1985

By J. E. Dutrizac, J. B. Sunstrum

"Lead was among the first metals known to man, and there are indications that lead was produced by the reductive smelting of galena in the mountains surrounding the Black Sea as early as 3000 B. C .(1). The early metallurgists soon discovered that silver could be recovered from lead bullion by cupellation techniques, and that the litharge product of the cupellation process could be reduced to give silver-free lead metal. By the Graeco-Roman era, lead-silver ores were being extensively exploited for both their silver and lead contents. Although silver was always of prime concern, a number of industrial uses existed for lead that included piping and sheeting. Subsequently, lead metallurgical practice developed steadily in different parts of the world, and was a relatively mature technology by the time of the European colonization of North America. In the United States, lead was produced on a small scale as early as 1720, and by 1800, the extensive Pb-Zn deposits of Missouri-Kansas and the upper Mississippi Valley were being developed(2). Many of the early American lead ores were low in silver, and seemed to have been worked primarily for their lead contents. Major uses of lead at the time included bullets, pipe and sheeting. During the period 1860-1890, the silver-rich galena ores of the western United States were exploited, and a number of lead smelters were operated. Although all the smelters produced lead metal, the recuperation of the associated silver and gold values was generally of prime concern. In Canada, lead deposits are abundant, and galenarich outcrops frequently attracted the attention of the early explorers and prospectors. In fact, the first record of a Canadian ore deposit seems to have been the observation of an argentiferous galena vein near Lake Temiscamingue in 17 41(3). The presence of silver-bearing lead deposits naturally attracted the attention of the mining community, and many outcrops were worked throughout the country. Because of the problems associated with transporting lead concentrates to existing American or overseas smelters, attention soon shifted to the establishment of local lead smelting facilities. The purpose of this article is to examine those early smelting efforts and to try to identify the first ""successful"" Canadian lead smelter."

Jan 1, 1999

By T. L. Tanton

Introduction Sturgeon River area, on the east side of lake Nipigon, Ontario, was the scene of a gold rush in 1934. According to press reports, the active staking during the latter part of the summer spread out rapidly from the early holdings until it embraced practically all of the lower basin of Sturgeon river in a belt up to ten miles wide extending west from Kinghorn for 35 miles to lake Nipigon. And, before winter set in, active prospecting and claim staking reached beyond lake Nipigon to areas where belts of ancient schists were known, notably west of Obonga lake. The writer&apos;s familiarity in past years with the areas to which attention has been directed admits of some observations on means of access in the field, the general geology, and the progress that has been made in identifying formations favourable for the occurrence of gold. Canoe Routes From the writer&apos;s observations in Sturgeon River area in 1917 and in the Obonga Lake area in 1933, it is recalled that there is, locally, a striking dif~ ference in the accessibility of the two areas by travel along canoe routes. Sturgeon River area is easily accessible from Kinghorn or Jellicoe stations on the Canadian National railways. Many small lakes, commonly of the linear type and with elongations east and west, occur among the forested rocky ridges near Jellicoe (Figure 1) and to the east and west along, or adjacent to, Sturgeon river. This stream has long stretches with gentle current between the falls and rapids. The drainage system provides excellent canoe routes (Figure 2), and the interruptions to travel (Figure 3) are remembered .as scenic features that were passed by portages with firm footing. West of Obonga lake, the streams flowing easterly to lake Nipigon may be followed through scenic canyons to a plateau of law relief characterized by extensive muskegs and shallow, brown-water lakes; and along small streams, where beaver dams have been destroyed, there are muck ponds where the usual alternatives of paddling or portaging can be dismissed in favour of the &apos;moosing&apos; method of progress with the canoe (Figure 4).

Jan 1, 1935

By J. C. O'Rourke

"Gibraltar Mines Ltd. has been operating its open pit copper mine for fifteen years. In that time over 140 million tons of waste and low grade material has been placed on the mine waste dumps. Laboratory and in-situ testing indicated that enough leachable copper was recoverable from the dumps to justify a leaching operation which would include the construction of a solvent extraction-electrowinning plant. The plant, which was commissioned in October, 1984 was designed to process 3,000 usgpm of dilute copper solution to produce 5,000 tons per year of high quality cathode copper.This paper deals with the start up and initial year of operation of Gibraltar's SX-EW plant. The three areas: leaching solvent extraction and electrowinning are presented with emphasis on the innovations which were made to help overcome the problems unique to die Gibraltar operation.THE LEACHING CIRCUITThe mines waste dumps are leached to produce a pregnant copper solution as feed to the SX-EW plant. To achieve this the dumps are sprinkled with a mildly acidic solution which percolates through the dump. In travelling through the dumps the solution contacts the surfaces of the liberated copper minerals and disolves the soluble copper into the form of copper sulphate. The leaching reaction is achieved by a combination of bacterial action to oxidize the suphides and solublization of the copper into the acid solution.For the bacterial action to occur the dumps must be at or near the ideal pH of 1.9 - 2.1 and the necessary elements of the reaction must be present. Because a bacterial reaction is involved there is a lag time between the time when the conditions for bacterial growth are achieved and when the bacteria are in their active growth phase. Laboratory tests with Gibraltar ore indicated that bacterial action takes two to three months to reach its maximum rate. See figure 1.The expected lag time for the planned rate of bacterial leaching prompted Gibraltar to start irrigating the dumps in late June 1986 for a plant start up in early fall. The lag time proved to be longer than expected and the dumps took until November to start producing copper at the expected rate. See figure 2."

Jan 1, 1988

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