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By Ian R. Morrison

"Abstract - The Izok Zn-Cu-Pb-Ag volcanogenic massive sulfide deposit is located 265 km south of Kugluktuk (Coppermine), Nunavut Territory, Canada, in the west-central Slave structural province. The Izok deposit is owned 100% by Inmet Mining Corporation and represents one of the largest undeveloped zinc-copper resources in North America. The Izok deposit is hosted within and near the top of a thick sequence of predominantly felsic pyroclastic rocks of late Archean age. The hanging-wall stratigraphy includes additional felsic volcaniclastic rocks, andesitic and basaltic flows, thin sulfide-rich iron formations, and turbiditic sedimentary rocks. The felsic volcanic rocks are intruded by intermediate dikes and sills, followed by gabbroic dikes and sills. Both intrusive suites are interpreted to be volcanic feeders to the overlying flows. All lithologies are subsequently cut by younger, irregular, granitic pegmatite and diabase dikes. The volcanic and sedimentary rocks are regionally metamorphosed to pyroxene hornfels grade. The massive sulfides occur within a large (kilometer-scale) Na-depleted sericitic alteration zone. The immediate footwall and hanging-wall rocks to the deposit are characterized by zones of muscovite- biotite-sillimanite, lesser chlorite-biotite-cordierite, and locally intense silicification and sodium metasomatism. All lithologies have been affected by younger Ca-metasomatism. As currently defined, the Izok deposit comprises a cluster of five complexly zoned composite lenses: the Northwest, North, Central West, Central East, and Inukshuk lenses. The first four lenses are amenable to open-pit mining, whereas the Inukshuk lens will require underground development. The total indicated mineral resource presently stands at 16.5 million tonnes with a grade of 2.2% Cu, 11.4% Zn, 1.1% Pb, and 60 g/t Ag. Inmet Mining Corporation is presently reviewing the economic feasibility of developing the property."

Jan 1, 2004

By Jom Kresta

Fundamental changes in the Bitumen Production process at Syncrude have led to the integration of Mining and Extraction operations. Successful operation of the integrated process requires the coordinated effort of several business units. Hence, decision-making has become increasingly more complex and better tools are required. IO2 is a collaborative effort to develop a proof-of-concept prototype of a distributed decision-support system for integrated short-term production and maintenance planning with fleet allocation and dispatching. The prototype includes three main modules (planning, dispatch, and maintenance), a shared database, and a simulation module. At this stage, this prototype is used to plan and operate a simulated mining operation. It is used to demonstrate the viability of the chosen systems architecture and algorithms, as well as the information flows and interactions necessary for improving the overall performance of the integrated operation. While the problem is demonstrated using Syncrude?s site, the methodology is applicable to most integrated resource allocation problems.

May 1, 2002

By G. H. Gale, M. A. F. Fedikow, L. B. Dabek

Rare earth element (REE) data are reported for ores, sul?de-bearing rocks and host rocks from massive sul?de deposits and barren sul?de facies iron formations in Proterozoic volcanic rocks of the Flin Flon-Snow Lake area, Manitoba. Some of the rare earth elements (e.g., Eu), are selectively leached in zones of intense alteration and redeposited together with the sul?des and silicatesin the uppermost parts of the alteration zones and at the rock-water interface. Chondrite normalized REE pro?les of layered sul?de ores and exhalites proximal to the deposits typically have positive europium anomalies and are enriched in light REE. In contrast, barren sul?de facies iron formations exhibit negative or ‘normal’ Eu pro?les. REE data are useful to explorationists in determining the presence of ‘off-hole’ metal-bearing sul?de zones, in distinguishing metal-poor sul?dic layers related to ‘economic’ hydrothermal activity from barren sul?de facies iron formation and in some instances as a means of establishing ‘ore-equivalent’ strata.

Jan 1, 1997

By G. H. Gale, M. A. F. Fedikow, L. B. Dabek

Abstract-Rare earth element (REE) data are reported for ores, sulfide-bearing rocks and host rocks from massive sulfide deposits and barren sulfide facies iron formations in Proterozoic volcanic rocks of the Flin Flon-Snow Lake area, Manitoba. Some of the rare earth elements (e.g., Eu), are selectively leached in zones of intense alteration and redeposited together with the sulfides and silicates in the uppermost parts of the alteration zones and at the rock-water interface. Chondrite normalized REE profiles of layered sulfide ores and exhalites proximal to the deposits typically have positive europium anomalies and are enriched in light REE. In contrast, barren sulfide facies iron formations exhibit negative or ‘normal’ Eu profiles. REE data are useful to explorationists in determining the presence of ‘off-hole’ metal-bearing sulfide zones, in distinguishing metal-poor sulfidic layers related to ‘economic’ hydrothermal activity from barren sulfide facies iron formation and in some instances as a means of establishing ‘ore-equivalent’ strata.

Jan 1, 1997

By L. M. Pidgeon

Abstract The production of metallic calcium, by reduction of lime with aluminium, was investigated. Using a modified form of the effusion method, the vapour pressure of calcium above the reaction: 4CaO + 2Al ~ 3Ca + CaO.AI203 was determined and found to vary from 1.0 to 1.3 mm. Hg from 1,150° to l,200°C. From the magnitudes of the various thermodynamic quantities involved, together with a consideration of the known chemical characteristics of the reaction, it was suggested that the reaction proceeded by reduction of lime with aluminium vapour. INTRODUCTION The conventional production of metals by reduction of their ores by carbon (or carbon monoxide) may be represented by the equilibrium: MO+ R [] RO+ M . . . . . . . . . .(1) where MO represents the metal oxide, and R the reducing agent (carbon or carbon monoxide). The free energy change ?F° associated with the reaction is given by -?F° = RTlnKp . . . . . . . . ... . (2) where R is the gas constant, T the absolute temperature, and KP the equilibrium constant. It should be noted that ?F° represents ? the standard free energy gained by the system and hence -?F° is the free energy gained by the surroundings - and that as ?F° increases K must decrease, and conversely. For metals above manganese in the electromotive series, the ?F° associated with reaction would quantity and from it can be seen that this corresponds to a small fraction for K. The equilibrium pressure of the oxide of carbon represented by RO is then too near its concentration in the furnace gases for the reaction to proceed in an internally fired furnace. For the practical reduction of such metals, fuel fired furnaces are inadequate. Electric furnaces may be used to give higher temperatures with lower concentration of carbon oxides and the reaction may still proceed. However, for many metals such as calcium and aluminium the formation of stable carbides or acetylides may prevent production of the metal by this reaction. In this case, either electrolytic methods must be used, orrecourse must be had to reducing agents other than carbon. It is the latter method which is under consideration here.

Jan 1, 1948

By A. Sinha, N. K. Babar, H. R. Kalihari, A. K. Chakraborty, M. R. Saharan, C. P. N. Pathak

A case study is presented on the optimization of cable bolting patterns in cut-and-fill stopes at the Chikla Mine, MOIL, India. Rock mass characterization, using Bieniawski’s RMR and Barton’s Q, indicated that the stopes lie in the “no supports required” to “spot bolting required” support categories. In contrast, the current practice of cable bolting in a 2.0 m by 2.0 m grid pattern for stope widths less than 12 m displays different results. Numerical modelling results are in line with the rock mass characterization study. It can be affirmed, as well, that no cable is carrying rock loads greater than 5 t. This fact is derived from observational behaviour of the crown pillar of an experimental stope with open stoping operations and hanging cables in the cut-and-fill stopes of the mine. An increase in the cable bolting grid spacing is, hence, recommended for implementation by mine management, based on the study described in subsequent sections. It is also suggested that the mine adopt the open stoping method, provided results of a rock mechanics instrumentation program in an experimental stope with a cable bolting grid of 2.5 m square are favourable.

Jan 1, 2007

By K. Morin

"Silver and gold leaching by cyanidation is a well-known process used around the world. This reaction is often accompanied by the leaching of other species throughtout the cyanidation circuit that causes the diversion of cyanide towards non valuable by-products (SCN-, Cu(CN)x, Fe(CN)x, etc.). One of the detrimental impacts of the formation of these by-products is the precipitation of silver-cyanide complexes that occurs when cyanide concentration is too low. Without any extra-addition of free cyanide, the precipitated silver is lost and cannot be recovered causing an important economic loss. To overcome silver precipitation, cyanidation circuits are often operated at higher cyanide concentrations. Recently, COREM’s research team has been working on a technology to measure and control silver precipitation during the cyanidation process. Moreover, a methodology has been elaborated in order to evaluate threshold cyanide level which corresponds to the cyanide concentration at which silver starts to precipitate. This technology has been tested on site on a real cyanidation circuit at Agnico Eagle’s LaRonde mine."

Jan 1, 2014

By Ilpo O. Auranen

Background: Mines of today are moving ever increasing tonnages of rock. The new ore bodies tend be of lower grade and located deeper in the earth?s crust. The industry?s focus to overcome this ?deeper but poorer? ore problem has traditionally been to lower the costs by increasing the scale of mining. This has resulted in an ever increasing size of mining equipment like drills, loading shovels, dump trucks. The main focus has been in reducing the cost per ton of mining. This cost focus has often resulted in increased waste rock dilution and more ore losses. From costs to value: Waste rock dilution means waste rock in the concentrator feed. It?s value in open pit mining is equal to the value of ore it replaces. In underground mining, waste rock dilution?s value is equal to the value of ore it replaces added to the costs of mining it. It is commonly accepted that waste rock dilution in open pit mining ranges between 5 and 20% and in underground mining between 5 and 40 % . In a typical large open pit mine ore value is from 20 to 60 USD per ton and mining costs are 1 to 3 USD per ton. One percentage point reduction in waste rock dilution represents savings corresponding to about 20% of total mining operation costs. The difficulty of reducing waste rock dilution comes from the complicated nature of the ores. Accurate ore waste-border detection has not been possible due to excessive costs. Core drilling combined with geochemical lab analysis is the most accurate method for locating ore waste borders. However, it is also expensive and time consuming, which is why it is often combined with somewhat lower cost methods like sampling drill cuttings from blasting drill holes or from reverse circulation drills and analyzing them in a laboratory. The main drawback with these methods, in addition high costs, lies in the sampling accuracy and in the quality control of the whole process when a large number of samples have to be handled manually. In every mine it is possible to halve waste rock dilution by using new on-line analysis technology .

Oct 1, 2009

By Brad Wells, Wayne Bull, Wayne Dickinson

"Antiscalants were evaluated for control of calcium carbonate (CaCO3) in gold processing applications. Under severely supersaturated conditions, 100 percent inhibition was achieved with 30 ppm antiscalant. Laboratory and field testing established that an antiscalant could both inhibit scale buildup and remove an estimated 0.5 kg/m2 per annum of pre-existing scale. A program was developed to control CaCO3 at a gold mill experiencing severe scale buildup. Annual accumulation of as much as 3 cm of scale in process piping was reduced to unmeasureable levels when 10 ppm antiscalant was applied. No scale-related upsets in mill operation occurred during a five-month trial period.INTRODUCTION Inorganic mineral scale is a crystalline deposit that forms when dissolved mineral constituents exceed the solubility of the respective mineral phase. Formation of mineral scale adversely affects a wide range of industries and is of particular interest in mining and metals processing. Scale affects numerous stages of metal extraction processes. Restricted flow in piping, nozzles, valves, screens, pumps and other components used to transport mineral slurries, metal concentrates, and tailings as well as scale in millwater and thickener overflow circuits are primary concerns (Alfano & Boffardi, 1994). Heap leaching, comminuation, flotation, vat leaching, carbon adsorption, and heat exchangers as well as oxidation steps used in processing sulfidic ores are among key unit operations that may be adversely affected (Simmons & Gathje, 2003). Antiscalant chemicals are designed to prevent formation and growth of mineral scales such as CaCO3. These chemicals, which include two primary classes – phosphonates and anionic carboxylate polymers and copolymers - are added to the gold processing stream at low parts-per-million (ppm) concentrations to interfere with nucleation and growth of scale forming crystals and to disperse fine particulate material. Application points may be located throughout the gold extraction process. Evaluation of the performance of antiscalants plays a critical role in determining the best product for a specific application. Performance evaluations include laboratory testing, pilot scale studies, in-plant trials, and ongoing performance monitoring during mill operation."

Jan 1, 2014

By J. -F. Lessard

Most underground mines rely on ore pass systems to ensure the transport of ore and waste. The design and operation of ore pass systems often rely on past experience, which is often anecdotal and at times contradictory in nature. This paper summarizes the current practice in Quebec underground metaliferous mines. The information was collected during a series of site visits in ten underground mines. The database includes information on ore pass configuration, support, materials transited, and overall ore pass performance.

Jan 1, 2006

By H. J. H. Pieterse

Agitation requirements in oxidation autoclaves can vary widely depending on the mineralogy of the feed material and the specific process selected. Before selecting an agitation system it is important that the client specify the desired throughput, mineralogy, chemistry, kinetics, compartments, overall retention time, and the desired oxygen utilization. In selecting the agitation system, the factors mentioned above as well as oxygen solubility, oxygen mass transfer coefficients, and specific power are evaluated. Much research effort has been directed toward open tank systems for fermentation and oxidation processes. However, for high pressure autoclave oxidation processes little information is available. In this review the author's experience in the autoclave industry and some of the available literature are utilized to compare the installed power per unit volume with some of the methods proposed for open tank oxidation. Equations that appear to give results close to that experienced in the industry have been identified. Further work is required in this area to develop a higher degree of confidence in the ability to predict the power required for any given case.

Jan 1, 2004

By C. M. Diaz, P. J. Mackey

Since its inception, in 1987, the Copper-Cobre series of conferences has established itself as the prime world forum for discussing advances in copper process metallurgy, in particular copper sulfide smelting, copper recovery by leaching and electrowinning and copper electrorefining. All copper producing countries have actively participated in the pyrometallurgy symposia held in the framework of these conferences. In 1999 and 2003, the close to 100 pyrometallurgy papers that were presented had to be accommodated in two volumes. In the present paper, the authors firstly provide a brief historical sketch of the Copper-Cobre conferences and then focus on providing a retrospective look at the dramatic changes that have taken place in copper smelting in the last few decades, primarily as discussed in the Copper-Cobre and precursor conferences pyrometallurgy proceedings. Topics include: the development of autogenous, increasingly more intensive, environmentally sound and energy efficient smelting processes; significant improvements in Peirce-Smith converting practice and at the same time the commercialization of alternative continuous converting routes; the consequent expansion and modernization of smelters; the increasing compliance with strict government environmental regulations in most copper producing regions of the world; the realignment of the copper smelting industry throughout the world; and trends in copper smelting R&D. A look at the future closes these reflections on the contribution of the Copper-Cobre conferences to the advancement of copper smelting technology.

Jan 1, 2007

By James W. Cassie

The Snap Lake Diamond Project is located approximately 220 km northeast of Yellowknife, NT, Canada at an approximate latitude of 63?35?. The project location is within the zone of continuous permafrost; measured ground temperatures are about ?2?C at 10 to 20 m depth below surface. The site is located on a peninsula jutting into Snap Lake. The site is generally barren of vegetation, except for some isolated areas with low trees located in topographic lows and near the shoreline of Snap Lake. As part of phased evaluation process with regards to the mining project, an Advanced Exploration Program was undertaken in 2000 and 2001. This program included the underground extraction of a bulk sample of kimberlite, which was processed on site. A Processed Kimberlite Containment (PKC) area was required as part of this program. The area chosen as the PKC facility was a small inland lake, called Inland Lake #1, on the peninsula. Construction of the PKC area required the construction of two dams, one on either end of the lake. At the south end of the lake, Dam 1 was constructed partially on a talik and partially on continuous permafrost. Dam 2 was constructed entirely on permafrost. The effects of the remote location, selection of suitable construction materials and permafrost conditions on the designs of the dams are discussed. A description of the design elements and construction schedule are provided. A review of the performance of these two dams, in terms of settlement, deformations, pore pressures and thermal regime, over the last several years is provided.

May 1, 2003

By L. Newman, R. Wyrick, R. Berezowsky

The Stillwater Mining Company operates an integrated mines, mill, smelter and base metals refinery complex in Montana, USA for the production of a high-grade platinum group metals (PGM) bearing concentrate, nickel sulphate crystal and cathode copper from the PGM-containing nickel-copper converter matte. The complex was constructed in stages over several years, starting with the mine and the mill, followed by the smelter and finally, a base metals refinery, which was commissioned in 1996. Due to the relatively small quantity of matte production, a straightforward refinery process was initially selected to extract the base metals, which reported to a mixed nickel-copper solution for further treatment off-site, and yielded a high-grade PGM concentrate suitable for treatment in a PGM refinery elsewhere. Subsequently, as a result of increased matte production, the process was modified to separate the nickel and copper, and to produce a nickel sulphate salt and cathode copper. This paper describes the design and operation of the original and the modified processes utilized in the base metals refinery (BMR).

Jan 1, 2004

By Staff

Introduction As the Sherritt Gordon property is still in the development stage, . it is rather difficult to write a proper account of the mining methods and milling practice that will be used. While the general scheme of operation will remain the same, there will undoubtedly be many minor changes to meet special conditions which we cannot forsee at the present time. Most of the work accomplished to date has followed the usual routine of preliminary development in the Canadian North. However, we have learned a few things which might be of use to others operating under similar conditions. The most important of these is the saving in power cost made possible by the use of Diesels as against steam power when using wood fuel, and the movement of freight by tractors when rail transportation is not available. Location The property i6 located about 90 miles north of The Pas, or 40 miles northeast of the Flin Flon mine. A branch line from Cranberry Portage, on the?C. N. railway, between Flin Flon and The Pas, connects the property with the outside.

Jan 1, 1930

By D. Pacey

"This paper reviews the installation, start-up, and operation of a comprehensive monitoring and targeting management technique in which all plant utilities (i.e., electricity, natural gas, compressed air, and water) are managed as controllable resources in the same way that raw materials, mined ore, personnel and capital are managed.The essential elements of monitoring and targeting (M&T) include:•gathering in a systematic, disciplined way the utility information by energy account centres;•closely monitoring the utilities used in each centre and comparing them with production on a common time base (e.g., kWh/ton per day);•bench marking and target setting to pro-vide attainable energy reduction goals which can be further refined as they are achieved; and•establishing an implementation plant that may provide efficiency enhancing benefits to various operating areas within the plant.The M&T system will reduce utility consumption per unit of production on a continuing basis.This would translate into sustained operating cost and waste. M&T will also provide better information for assessing the plant operation and progressively implementing further enhancements."

Jan 1, 1997

By W. E. Roscoe, H. Agnerian

The contour method can be used to estimate the mineral resources of a mineralized body. It relies on geological interpretation and allows the geologist to determine the plunge of a mineralized ore shoot along the dipping plane of the mineralized body. The technique is best applied to estimate tonnage and average grade of relatively planar bodies, i.e., where the two dimensions of the mineralized body are much greater than the third dimension, such as Archean lode gold vein deposits or mantos of massive sulphide mineralization, based on exploration drilling. For these types of deposits, the contour method can provide a clear view of the “mineralization landscape” with “peaks” and “valleys” along the plane of the mineralized vein. The technique also reduces the “nugget effect” of local high values commonly present in gold deposits. With sufficient drill hole information, the contour method may be used to estimate the tonnage and average grade of individual blocks within a mineral deposit. General procedures for the contour method are described in this paper, along with examples and explanatory notes.

Jan 1, 2002

By J. L. Jambor, E. J. Reardon, E. C. Appleyard, J. A. Cherry

Abstract- Since the 1957 initiation of deposition of mill tailings at the Heath Steele mine, New Brunswick, periodic detailed studies of the tailings geochemistry and mineralogy have been con­ ducted, thereby providing a unique opportunity to examine the geochernical evolution of a sulfide- rich mine-tailings impoundment. When deposited, the tailings contained about 85 wt% sulfide minerals, 1 wt% to 3 wt% carbonate minerals, and 12 wt% to 14 wt% aluminosilicate minerals. Mineralogical examination of samples of the tailings from deep in the impoundment indicates that little oxidation occurred prior to tailings deposition. The mill process-water deposited with the tailings was neutral to slightly basic and contained low concentrations of Cu, Pb, Zn and As. Sampling of the tailings five years after deposition indicated that sulfide-oxidation reactions, occurring near the tailings surface, had decreased the tailings-water pH and increased the concentrations of Fe, SOi-, Cu, Zn and Pb (up to 10 000 mg/L, 25 000 mg/L, 2.5 mg/L, 200 mg/L, and 4.5 mg/L, respectively) . In 1975, ten years after decommissioning, the tailings could be divided into three zones: an upper oxidation zone in which sulfide oxidation is actively occurring, a hardpan layer, 15 cm to 20 cm thick, composed of tailings cemented by secondary Fe minerals, and an underlying reduction zone in which the tailings are relatively tinoxidized (Boorman and Watson, 1976). Chemical analyses indicated the presence of low-pH waters containing very high concentrations of SO42- , Fe, and other metals (up to 63 600 mg/L SO42-, 28 600 mg/L Fe, 10 mg/L Pb, 600 mg/L Cu and 4000 mg/L Zn), suggesting that there was a period of intense sulfide oxidation between five and ten years after tailings deposition. More recent sampling, conducted some 20 years after the end of tailings deposition, indicates that relatively little change has occurred in the last ten years, thereby suggesting that the hardpan layer has provided an impediment to the downward movement of dissolved constituents . Analyses of the tailings solids show increased accumulations of many metals at or near the depth of the hardpan layer. Evaluations of the long-term evolution of the tailings can be enhanced by comparison of descriptions of the tailings with descriptions of the gossan that overlay the massive­ sulfide orebody prior to mining. At Heath Steele, this comparison suggests that many of the mechanisms controlling the distribution of metals in the tailings will probably persist over geologic time.

Jan 1, 1992

By W. D. Sinclair

"Between 1934 and 1956, nearly 31 million grams of gold were produced in the Matachewan area from deposits on properties belonging to Young-Davidson Mines Limited and Matachewan Consolidated Mines Limited. In 1979, production from these properties was renewed by Pamour Porcupine Mines Limited.Most of the gold produced has come from relatively lowgrade deposits (3.4 gl t or less) that occur in an east-trending belt of dykes and irregular bodies of trachytic syenite. Although small by comparison, these deposits have some of the characteristics of porphyry copper-molybdenum deposits. The mineralized syenite contains up to 5 per cent disseminated pyrite and is highly fractured and cut by quartz and quartzcarbonate veinlets. Gold occurs as native gold along fractures, in quartz veinlets and as minute inclusions in pyrite. Minor chalcopyrite and trace amounts of scheelite, molybdenite and galena are also present. Mineralization was accompanied by potassic alteration, consisting mainly of potassium feldspar.Some gold was also produced from small, but higher-grade, deposits in volcanic rocks. These deposits occur near the syenite-hosted deposits and consist of quartz veins and stringers mineralized .with pyrite and gold. Minor gold is present in some low-grade, porphyry-type copper-molybdenum occurrences and, locally, in large, isolated quartz veins.Lead isotope analyses suggest that the gold deposits were formed at about 2700 Ma and were locally remobilized at approximately 2400 Ma, following the Kenoran orogeny. The deposits appear to be genetically related to felsic magmas emplaced in an epizonal environment and deposited from hydrothermal solutions relatively concentrated in S and K20."

Jan 1, 1982

By Guy Chevalier, Jean-Guy Chartrand, Louis Barrette, Daniel Michaud, Michel Fortin

"COREM has developed, specifically for the mineral industry, a mobile laboratory using the Laser-Induced Breakdown Spectroscopy (LIBS) technique. It has been used to perform onstream analysis (OSA) of the chemical composition of slurries in both gold-ore and iron-ore processing plants. Industrial tests have been carried out respectively at La Mine Doyon (Cambior), a gold-ore processing plant near Rouyn-Noranda, Quebec, and at QCM's iron-ore pelletizing plant in Port-Cartier, Quebec. The aim was to demonstrate that the LIBS technique is capable of providing information that meets the needs for process control. C, Si, Ca, Mg, Mn and Al were the elements targeted on iron ore while Cu, Fe, Ag and Au were targeted for gold ore. The results obtained meet or exceed the requirements for process control and consequently clearly show that the LIBS technique is suitable for industrial applications.The feasibility of OSA, through in-plant validations using COREM’s LIBS mobile laboratory, has been demonstrated in four essential aspects: slurry sampling, LIBS chemical measurement, robustness for continuous operation, and functionality for industrial operators.INTRODUCTIONThere is a general need for optimization in the mineral industry. Process control is a favoured means to achieve this goal (McKee 1991). Real-time knowledge of process status is the key to effective control (Da Silva 1994). Only a few techniques, such as XRF, NMR and digital imaging, are currently amenable to chemically monitor industrial processes (Braden 2002). LIBS is an emerging technique with appealing characteristics for on-line and real-time measurement of bulk material (Lorenzen 1992, Palanco 2000). While our ultimate goal is to provide our industrial partners with performing monitoring tools, a shorter-viewed objective is to demonstrate the feasibility of developing a LIBS system for OSA. The LIBS technique has shown interesting capabilities for measuring solids, liquids, dusts and gases. Our work has focused on mineral slurries (Barrette 1999 and 2001, Turmel 1999). Special equipment is needed to perform in-plant validations. A mobile laboratory based on the LIBS technology has been assembled for this task."

Jan 1, 2004