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? A BARGE WILL BE INSTALLED IN THE BOTTOM OF A COMPLETED PIT. ? THE PROVEN BAUER CUTTER TECHNOLOGY WILL BE INSTALLED ON THE BARGE AND A DEWATERING PLANT WILL BE INSTALLED NEAR THE BASE OF THE PIT RAMP. ? THE WATER LEVEL IN THE PIT WILL BE RAISED BY 20- 40 METRES AND REMAIN CONSTANT OVER DURATION OF MINING. ? THE CUTTER WILL BE LOWERED TO MINE 10 METRE DEEP BENCHES ON A CHECKERBOARD PATTERN TO AN ULTIMATE DEPTH OF 200 METRES. ? THE 5? MINUS CUTTINGS WILL BE PUMPED TO THE DEWATERING PLANT, DEWATERED, AND TRUCKED TO SURFACE.

May 1, 2003

By Pierre Boulanger

This paper describes a generic technology for modeling in real time the various operations of surface mines. Based on various sensor fusion techniques and high-speed network, a virtual reality model of a mine capable of representing not only static infrastructures but also mobile equipment such as trucks and excavators is created. The modeling technique deal with a wide range of sensors such as survey data, satellite pictures, ore concentration, machine condition indicators, video, sound, real-time GPS locator, range sensor, etc. The model created allows for every piece of information to be displayed in context allowing for a better temporal and spatial correlation of events by mine operators, planners, and supervisors. Syncrude, a Canadian mining company located in northern Alberta, is planning to use this technology, not only to understand oil sands mining processes better, but also to give decision-makers better access to the timely, accurate information from the field that they need to make informed decisions.

May 1, 2001

By Zygmunt Kowalczyk

"A theory permitting the determination of expected surface deformations for planned underground exploitation is proposed and discussed in this paper. The relations between surface subsidence and various methods of mining are analyzed, and allowable horizontal .deformations for protected objects, according to the degree of their significance, are reviewed."

Jan 1, 1966

By Y. H. Hatzor, D. Bakun Mazor

"A new model for thermally-induced block displacement in discontinuous rock slopes is being explored inside a Controlled Climate Laboratory (CCL). The model is based on a hypothesized wedging mechanism that takes place inside the filled tension crack of a discrete block that rests on an inclined plane. Irreversible block sliding is assumed to develop in response to climatic thermal fluctuations and consequent contraction and expansion of the sliding block material. While an analytical expression for this model is already available, we are exploring the possibility of obtaining such a permanent, thermallyinduced, rock block displacement, under fully controlled conditions in the laboratory, to enable effective analysis of the sensitivity of the mechanism to geometry, mechanical properties, and temperature fluctuations.A large scale (approximately 0.5 m3 in volume) model consisting of four blocks has been installed in the CCL. The CCL permits accurate control of atmospheric temperature (?0.5 °C) from 6 °C to 45 °C. Harmonic temperature function is planned in advance and is controlled by a computer program throughout the experiment process. The minimum length of the temperature cyclic period for thermally-induced block displacement is calculated here based on the existing analytical model.Due to the fact that this experiment is still in progress, preliminary results will be available in the next few months. The obtained results may explain sudden triggering of rock slope failures that are not correlated with common failure triggers such as seismic activities, pore pressure build-up, or freezing and thawing of water inside rock joints."

Jan 1, 2015

By J. W. Keyes

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

Jan 1, 1964

By H. Kroog

"This contribution reports recent mineral sorting experience and lists the possibilities of the newest generation of automatic mineral sorters with a special focus on electromagnetic detection, optoelectronic technology and the latest innovative X- Ray transmission units. In industry, mineral and ore are still hand-picked and current automation potential is only utilised in a handful of specific applications. Recently a number of technological developments took place that may increase sorting potential, save mill costs and lead to lower overall tonnage costs. The mentioned technologies provide automatic dry separation of minerals and rock as an alternative for conventional processing. The MikroSort technology is environmentally friendly:•,The remaining waste is free of contaminations•,The dump volume is reduced•,The concentration of the high grade material•,at an early stage saves resources"

Jan 1, 2005

By Nelson Hogg

"THE NOR-ACME gold mine is located close to the north shore of Snow Lake, at 54°53' north latitude and 100°03' west longitude. It is about 75 air miles east of Flin Flon, and 36 miles by road from Wekusko at Mile 81 from The Pas on the Hudson Bay Railway.The claims which contain the Nor-Acme orebodies were staked in 1927 by C. R. Parres, who, with his sons, held the ground until 1938 when Nor-Acme Mines Limited was incorporated. In 194 l, on the recommendation of F. Ebbutt (1), this property was optioned by the Howe Sound Exploration Company, Ltd. A program of 50,000 feet of diamond drilling resulted in the outlining of two orebodies containing an estimated 5,000,000 tons with an average grade of 0.15 ounces per ton to a depth of 1,000 feet.A period of inactivity during the war years was followed by the construction of a mining plant capable of treating 2,000 tons of ore per day. Milling started in March 1949, and to the end of 1954 has resulted in the production of 383,793 ounces of gold and 32,906 ounces of silver from 3,331,403 tons of ore.General GeologyThe general geology of the Snow Lake area is described by J. M. Harrison (2), and will not be discussed here except insofar as it affects the Nor-Acme orebodies. The orebodies occur in rocks of Precambrian age that have been classified by Harrison as the Snow group, comprising meta-morphosed sediments interbedded with acid and basic volcanic rocks. The volcanic rocks include pyroclastic types which are unique in the region, a fact which led to their classification in the Snow group rather than the more widespread Missi group that generally occupies the same stratigraphic position above the Amisk group (Figure 1)."

Jan 1, 1949

By M. Oliazadeh

The bioleaching of fine low grade copper sulfide ore is described. A sample of ground (> 1 mm) copper ore from Sarcheshmeh, Iran, was subjected to a series of agglomeration procedures and subsequently leached. The best agglomerate was obtained by using lime and sulfuric acid as binders. A column bioleaching test of agglomerated ore was then run for 105 days along with a non-inoculated control column having thymol as bactericide. The results show that about 77.5% of copper was extracted in the bioleaching, whereas the control column achieved only 57% copper leaching.

Jan 1, 2007

By Patty Simpson, Sarah Benson

In an effort to improve consistency across the nuclear industry, over the past 10 years the Canadian Standards Association has developed a series of standards related to environmental management at Class 1 nuclear facilities and uranium mines and mills. Canadian uranium mines have been required by the Canadian Nuclear Safety Commission to align their existing monitoring programs with these standards. Orano’s McClean Lake Operation’s monitoring program, which was originally presented in the Environmental Impact Statement in 1991, has evolved through regulatory change, new scientific understanding, additional projects, and experience into a comprehensive set of monitoring requirements for effluent, groundwater, air emissions, and the receiving environment. The exercise of aligning the existing practices with the standards provided an opportunity to review and rationalize the entire program. To document the process the original “Environmental Monitoring Program Design Document” was restructured and developed as both a repository for historical information and monitoring rationalization, and also to be a living document describing on-going studies and supplementary work that are outside of the regular environmental monitoring program and may justify future changes to the program.

Jan 1, 2020

By A. Gutiérrez

At present, most ore milling is done in tumbling mills, consisting of huge cylinders turning on their horizontal axis. This rotational motion causes the ore charge and the grinding balls to interact by impact, compression and abrasion, to reduce the size of the ore. This method is very inefficient since only around 12% of the energy supplied is used for actual grinding, the rest being used in moving the large mass of balls and ore within the mill. This work aimed at better understanding the milling process by evaluating the movement of the load through the Discrete Elements Method (DEM). First, the number of bodies (balls and ore) moving inside the mill is determined, to correlate to the potential impacts that constitute one of the milling mechanisms. Then, the number of high energy impacts is evaluated to correlate to the bodies in motion. At this stage, a differentiation between grinding media and the ore, as well as between the types of contact taking place in the mill, has to be made. This latter issue is of great relevance since from the five identified contact types (ball-ball, ball-liner, ore-ore, ball-ore and ore-liner) only the last three actually contribute to the milling process, the other two causing structural damage to the mill.

Jan 1, 2007

By Yves Gagnon, Nicole Gauthier, Daniel Kelly, Michel Rocheleau

"The Larder Lake-Cadillac Tectonic Zone in Rouyn-Beauchastel area (a 20 km long segment) contains distinct lithotectonic units bounded by artastomosing or sub-parallel shear zones. The gold deposits in the area are located within highly strained portions of E-W striking shear zones. The ore-zones are structurally controlled disseminated and vein-type lodes. They are spatially related either to megascopic z-shaped folds which refold E-W shears, or to areas of intersection between E-W shear zones and late NE-striking faults. Several ore bodies are enclosed within ultramafic rock units or at their contacts with adjacent units. All the lenses are elongated down dip and distributed, both loterally and vertically, ""en echelon"" . The ore zones tend to decrease in size and number away from the hinge zone of the megascopic z-shaped folds.Two main styles of mineralization are recognized at the Astoria deposit and at for most of the deposits in the Rouyn-Beauchastel segments a discontinuous ""Carbonate Ore"" and a more continuous, albitic sulphide-bearing ""Flow Ore"" (terminology used at the Kerr Addison mine).The typical ""Carbonate Ore"" zones are found within strongly arbonatized blocks bounded by conjugate late NE or NNW-striking fralls. The carbonatized blocks display features characteristic of llrittle to brittle-ductile deformation developed in shear zone system. Vein systems in the ""Carbonate Ore"" occur within shear fractures(D R)tension fractures (T), or within brecciated rocks adjacent to NS llrilcing faults. The ""Carbonate Ore"" zones are displaced by numer IU lole E-W, NNE, NW and N-S faults.The ""Flow Ore"" zones are found within predominantly ductile portions of E-W shear zones. The sulphide mineralization is finely disseminated and regularly distributed within mylonitic foliation or associated with veins and veinlets sub-parallel or oblique to the mylonitic foliation, in which case the mineralization is coarser grained. The gold mineralization is a late Archean event postdating the formation of the penetrative fabric, and synchronous with the end of the dextral oblique-slip motion."

Jan 1, 1990

By J. A. Finch, M. Mimezami, M. S. Hashemi A.

The adsorption of magnesium and calcium ions on sphalerite and pyrite is discussed. These ions are common in process water and often considered be detrimental in sulphide flotation. The study attempts to characterize the adsorption mechanism, which has been less studied than the oxide system. Generally for oxides metal ion uptake occurs when cations are hydrolyzed. Our findings reveal that magnesium and calcium adsorb on sulphides at pH below hydrolysis, but adsorption is not chemical. At higher pH, charge reversal occurs that does correspond to presence of hydrolysis products. Advances

Jan 1, 2009

By N. M. Steven, J. M. Moore, J. L. Kuyper

Mineral exploration and research has led to the identification of four types of tungsten mineralization in the NE-trending, Iargelyensialic, inland branch of the late Proterozoic/early Paleozoic Pan-African Damara Orogen. The mineralization (mainly scheelite) occurs in a terrane (the Central Zone) that is characterized by multiple deformation, greenschist/amphibolite-facies metamorphism and numerous peraluminous S-type granitic and pegrnatitic intrusions. The four major types of scheelite mineralization are (in ascending stratigraphic order of host rock): (1) thin (5 cm to 2 meter) quartz+plagioclase+diopside ± scapolite calc-silicate granofels with fmely disseminated scheelite (scheelite-bearing granofels; SBG) developed as discontinuous intercalations within the calc-silicate rocks of the Khan, Spes Bona, Oberwasser and Kuiseb Formations (Nosib and Swakop Groups); (2) gamet+pyroxene± vesuvianite+scheelite ± fluorite replacement skarns hosted by Okawayo, Karibib and lower Kuiseb Formation marbles in the aureoles of late tectonic, unrnineralized leucogranites; (3) metaturbidite-hosted scheelite- (and possibly wolframite-) bearing tourmalinites, both stratiform and crosscutting, in the Kuiseb Formation located in the 'cleavage fronts' of late tectonic granitoids; (4) scheelite-bearing vesuvianite+quartz+garnet skarnoid rocks (several hundred meters in length) in the calcareous metaturbidites of the upper Kuiseb Formation. These scheelite occurrences were discovered by the follow-up of a large number of tungsten anomalies located during a regional stream sediment sampling program. Detailed mapping, ultra-violet surveys and rock sampling proved to be the most effective surface exploration techniques for locating mineralization; soil sampling and geophysics were of almost no use. The largest, continuous scheelite-bearing skarn of economic interest, Otjua, has a drill-indicated reserve of 250 000 tonnes at 0.53 wt5 WO3 and 8 wt% CaF2•

Jan 1, 1994

By Rachel S. Prudhomme, Michael L. Jeremic

"The consolidation of mine backfill by free zing is a concept which has been under study in the mining engineering laboratories at Laurentian University during the past three years. The proposed underground application of freezing involves the artificial freezing of in-situ backfill by means of a refrigerated brine solution which would be circulated through a circuit off freeze pipes. The most interesting findings to date are: (I) The uniaxial compressive strength off frozen fill is considerably higher than that of cemented fill in a ratio of 20:1 (aggregate to cement), and (2) load deformation characteristics of the frozen fill suggest that a higher ice content results in more brittle failure and that samples which have initially failed can still support up 10 one-half the ultimate load without additional failure.Although at this early stage a derailed economic analysis of the procedure has not been possible, a general evaluation of the costs involved suggests that the method could be competitive with the hydraulic placement of cemented backfill. Results obtained from laboratory investigations indicate that the freezing of backfill is a structurally feasible option .IntroductionGround consolidation by freezing is based on the principle of transformation o f water contained in an aggregate-slurry into ice, which act s as a binding agent between the bulk particles, and which results in the transformation of a granular soil into a ""rock -type"" material. The abrasive nature of the particles enhance s the strength o f the ice by providing internal friction. As a result of this internal friction, the chances of plastic flow of the frozen mass when in a loaded condition are reduced(1). T he strength of a frozen aggregate is al so dependent on the grain shape of the aggregate particles which may or may not promote particle interlocking(2) and on the temperature since, below the freezing point of the water/aggregate mixture, the strength of the frozen mass increase s with decreasing temperatures.In the shaft sinking industry, artificial freezing of loose , unconsolidated soils has been in practice for several decades. Shafts for Saskatchewan 's potash industry have been sun k to depths of 3000 feet by employing artificial freezing to stabilize the Blairmore Formation(3). The common procedure for freezing ground for shaft sinking is summarized as follows(4,5): I) 150 mm diameter bore holes spaced on 1.5 m centres are drilled in a circular pattern with a radius of 1.5 m outside the periphery of the proposed shaft."

Jan 1, 1985

By Chester J. Kuryliw

"The Geology of the Cochenour Willans mine, Red Lake district, was described by J. P. Nowlan in the Jubilee Volume. The complexity and intense alteration of the rocks obscured any clear relationship between ore and structure, and Nowlan recognized that future work might greatly modify the ideas then existing. This has proved to be the case. Subsequent detailed work has established the stratigraphic sequence in the Keewatin formations which are the host rocks of the orebodies, and has shown that the principal orebodies are confined to two adjoining bands of tuff.On this basis, it is now possible to map the complex structures with greatly increased accuracy. The amount of drilling and underground exploration required to find new ore has been much reduced, and indicated ore reserves are now substantially greater than at any time in the past. Daily production in 1955 averaged 187 tons per day at 0.578 ounces of gold per ton.The sequence of Precambrian events in the mine area is summarized in the following table:Intrusion of post-ore dykes.Ore formation.Intrusion of post-faulting, pre-ore dykes.Deformation: folding and shearing; alteration; faulting.Timiskaming sedimentation.Unconformity.Keewatin: volcanics and sediments, host rocks of the orebodies, described in detail in the next section.KeewatinFrom oldest to youngest, the Keewatin rocks in the mine area can be divided into six formations, which have been given local names:“Grey"" LavaA fine-grained dacitic to andesitic lava, amygdaloidal and locally containing well formed pillows. Certain areas are sheared and chloritized to a degree that makes identification difficult. The rocks are carbonatized throughout; the most persistent carbonate veins parallel the upper contact. Thickness more than 1,000 feet. (Minor mineralization, type 1 I -See section on Ore Formation)."

Jan 1, 1949

By R. C. Wallace

Introduction At a time when the mineral-bearing areas in Manitoba are being fi very keenly explored because of the major developments which are now taking place in that Province, it may be timely to give an analysis of the mineralization in the two major types of deposits, with a view particularly to assist in the intensive prospecting to which the mineral-bearing areas in the Province are now to be subjected. As is already well known, there are two principal types of mineral deposits in the territory- copper-zinc and gold; and attention will be confined almost exclusively in this paper to these two types, with some indication of such modifications of the main types as may be found to be of economic value. The details of the paper are drawn from what is yet a small part of the area of the Province available for prospecting. Enough is now known of the types of mineral deposits which are described in the following pages to justify the expectation that these types may serve as criteria in the search in the newer districts. Doubtless new types will also be found as investigations proceed.

Jan 1, 1928

By S. J. Buckler

Many metallurgical processes use submerged gas injection to increase the contact area between gas and liquid phases in order to promote faster overall reaction rates. The work presented in this paper is part of an experimental investigation to measure the gas-liquid interfacial area available in bubble plumes within molten metal baths. Measurements of the nitrogen absorption rate during submerged injection into molten iron baths have been used to estimate the interfacial areas of the bubble plumes. A theoretical model was used to predict the superficial nitrogen transfer rate under mixed liquid phase mass transfer -interfacial reaction control. The interfacial area was calculated by dividing the overall nitrogen absorption rate by the superficial rate. Reasonable agreement was obtained between the interfacial areas measured in our laboratory with data from a comparable system reported in the literature. The experimental results were compared with the predictions of three theoretical models and one empirical correlation. The models of Turkdogan (3) and Zhao and Themelis (5) gave the best agreement with our data.

Jan 1, 2004

By F. Ford, J. Dong

"Laboratory and miniplant testwork was carried out to reject uranium from an Iron Oxide Copper Gold (IOCG) deposit that contains an average of 60-70 ppm uranium with high variability. The dominate copper bearing mineral is chalcopyrite with minor bornite and trace covellite. The main gangue minerals are iron-rich chlorite, quartz and magnetite. The main uranium-containing mineral is uraninite (UO2). Downstream processes require the maximum uranium level of no more than 80 ppm in final copper concentrate.Mineralogical examination reveals that uraninite particles are finely disseminated with grain size of less than a few microns in the deposit requiring fine regrind in order to liberate uraninite from chalcopyrite particles. A lab flowsheet was developed to include ultrafine regrind of the rougher concentrate to the particle size of 80% passing 11-13 µm with up to four stages of recleaning in flotation columns. Finer regrind to 80% passing 5-6 µm may continue to improve liberation but does not decrease uranium level in the final concentrate. In each of the successive recleaning stages by flotation column after fine regrind, uraninite is rejected into the column tailings, with the final copper gold concentrate containing less than 70 ppm uranium from a feed containing 83 ppm uranium. The conceptual flowsheet was further tested in an integrated, continuous miniplant showing average 94% copper recovery with average 74 ppm uranium. The uraninite particles that report to the final concentrate are either locked with chalcopyrite or entrained as liberated particles of less than 5 µm in size."

Jan 1, 2016

By J. Salazar

This paper presents a dynamic simulator of the semi-autogenous grinding operation deduced from first principles (non-stationary population balances) coupled to a Model-Based Real-Time Optimization (MBRTO) system to study optimizing strategies for an industrial semi-autogenous mill using genetic algorithms. For the design of the MBRTO both a validated first-principle steady-state model adapted on-line and a dynamic model were used to study the transient effect. The objective function considered is to maximize the ore throughput or the fine-product rate by changing the power-draw. For these purposes, these two performance indexes were introduced as fitness functions in the genetic algorithms. The three decision variables used were the ore feed rate, the water feed rate and the mill rotating velocity. The optimization problem was solved using a classical genetic algorithm with constant population. Constraints in the power draw and the filling level were handled by assigning a penalty function to the fitness function. Simulation results using industrial data for a large 1800 t/h copper-ore mill show the effectiveness of the system in finding the optimum controller set-points as a function of the process-operating constraints; mainly maximum ore feed rate, power draw and filling level. Starting from non- optimum conditions, the optimizer is able to improve the global performance index by 50% relative to a base case.

Jan 1, 2007

By J. Suarez

The innovation initiatives for world-class mines today and in the future focus on automation and robotics in equipment and machinery. These innovations have benefited from the development of advanced tools in global positioning, the Internet, the capabilities of data storage, transmission and handling, and intelligent systems. Forward-thinking companies have been investing in automated trains and trucks, selfdrilling rigs, intelligent flotation systems, etc. Considering that the emphasis has been mostly laid on improving the performance and reliability of the equipment and machinery, it is not surprising that the visibility was geared toward mechanical, electrical, process, and metallurgical engineering. One of the critical disciplines, mining engineering, has had less shine from the spotlight of automation and robotics. Mining engineering is concerned with the design, the operation, the management, and the reclamation of mines and mining facilities. In essence, it does benefit from the advantages of automation and robotics in equipment and machinery. However, being ?left out? in the innovation process may impede the core capabilities of this engineering discipline, which would create a two-tier knowledge and technology system in mines. For example, shovels in an open pit may be equipped with technology that can send real-time data of actual payload, number of buckets, mechanical status, etc. But, are mining engineers aware of the possibility to incorporate the data into their planning tools in order to have updated forecasts and accurate information? And if they are, are they equipped with the tools to reap the rewards of those technologies? And if so, is the personnel trained to use the tools efficiently? This paper is a contribution to the discussion on the technological advances in mining engineering. For the authors, the roadmap to the future requires a smart integration of technological advances in hardware and software so that mining engineers are not only aware of the possibilities, but also trained and capable of using the available tools. Essentially, this paper believes that automation and robotics in equipment alone will be undervalued and underutilized if the owners and drivers of the mining systems (i.e. the mining engineers) do not possess the tools and training to integrate the technological advances into the functions of mining engineering. The intent of the paper is to analyze the available technologies for the core mining engineering functions of design, planning, scheduling and optimization. The analysis will lead to discussions and suggested solutions on integrative technology solutions that would create optimum value for the mining processes and equipment technologies.

Aug 1, 2013