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By Gaby Steinbach

A novel synthesis is described for the preparation of supported gold catalysts for low-temperature CO oxidation. The novel route is a variation of the inverse co-precipitation method and requires amphoteric metal hydroxides, with zinc and aluminium hydroxides chosen. Catalysts were prepared by inverse co-precipitation (ICPa) and by the novel route (ICPb) and tested for CO oxidation at 60ºC. Preliminary results showed the novel route to be superior to ICPa; the catalysts prepared by ICPb showed higher catalytic activities. The effect on catalytic activity of parameters such as pH, calcination temperature, washing of precursor and calcined catalysts and ageing was investigated. A preparation pH of 9 and a Zn to Al mass ratio of 3 proved to be optimal. It was possible to remove residual chloride by washing thereby enhancing catalyst performance. Ageing and Mg citrate addition were also shown to improve catalyst performance. Characterization of the catalysts revealed that better activities were obtained when the support was more crystalline and had a higher surface area. The presence of carbonate-containing species in the support decreased catalytic activity. 197Au Mössbauer Spectroscopy revealed the presence of Au3+ when catalysts were heat-treated at temperatures of 200ºC and below. Catalysts calcined at 400º had 75% of the gold present as Au0, with the remaining gold present as Au3+.

Oct 1, 2003

By John L. Rich

A New Field for the Mining Industry A new field for the mining industry is in prospect. This is the recovery of the petroleum which has been left in the ground in the older oil fields whose yield has declined until they are no longer profitable to operate in the ordinary way. The fact that an oil field has reached the economic limit for operation by pumping does not necessarily 'mean that the oil has ? been exhausted. In many such cases not over one-fifth, or even one-tenth, of the total oil content of the field may have been taken out when this stage has been reached. This means that there exists a large reserve of oil which can be recovered by whatever process best takes advantage of the natural conditions under which the oil occurs. Mine drainage is one of the processes by which this oil may be recovered. Its relative importance will depend partly on the ingenuity and wide-awakeness of mining engineers, and on mining capital. The opportunity is undoubtedly at hand. In order that we may be in a position to judge the possible advantages of mining methods for the recovery of this oil, I have reviewed, on the following pages, the methods now in use, and have discussed the physical conditions of the reservoir rocks in which the oil is found, and the suitability of the various recovery processes, in use or proposed, in the light of these reservoir conditions.

Jan 1, 1927

By A. Vervoort

The concepts of geostatistics are commonly difficult to understand for students, even for those with a strong mathematical background. A main problem is certainly the link between the variation of the parameter value in space or time, and the calculated experimental semivariogram as a function of the interdistance, but over the entire area. Another problem is to differentiate between directional and omnidirectional semivariograms and the way they have to be interpreted. Furthermore first time users of geostatistics often do not want to struggle through different books with the whole geostatistical background before using it. Too often this leads to using geostatistics as a black box with all the risks linked to it. That is the main reason that we developed some interactive tools based on the free public domain software R. We aim by developing three different modules to assist individual first time users, but also regular students and teachers to get a feeling of the importance of the various aspects in conducting a good geostatistical study. The modules show directly e.g. the impact of deleting, moving or adding data points, of changing parameters of the semivariogram model and of the geometrical parameters during kriging. The first module allows making the link between the variation of the parameter values in one direction and the calculated experimental semivariogram, and this for different given datasets. The second module covers two dimensional datasets and apart from calculating the semivariogram values, models can be fitted to the calculated data. A distinction can be made between omni-directional and directional semivariograms, including the effect of the tolerance on the angle and lag-distance. The third module focuses on ordinary kriging of point values in two dimensions. The aim of the interactive tools developed is limited to educational purposes only, i.e. to visualize some basic concepts for different pre-selected types of spatial variation. They do not intend to do a geostatistical study on a certain dataset (for this purpose there exists already a broad spectrum of good software packages). The original intention of developing the three modules was aimed at users with no or a limited amount of geostatistical experience. However, these modules seem also very useful to remain critical for experienced users. One should never forget that in a real application one does not have a complete view of the phenomenon studied. The geostatistical interactive modules can be accessed through: www.bwk.kuleuven.be/geostatistics . We compiled a manual to facilitate the use of the modules and to guide the user we formulated several questions to be solved for each module.

Aug 1, 2013

By V. N. Bryant

"1.0 SYNOPSISThe Colomac property, located 137 miles by air northwest of Yellowknife in Canada's NWT. will recover gold from 10,000 tons per day of open pit ore.The decision to proceed with the project was based on a mineable ore reserve of 16 million tons (including dilution) averaging 0.064 oz/ton. After an in-fill drilling program in the latter half of 1988, Colomac's mineable, proven and probable ore reserves were increased to 28 million tons averaging 0.056 oz/ton of gold.A feasibility study completed by Wright Engineers for Neptune Resources in December 1987 examined the possibility of vat leaching followed by C.l.P. and electrowinning. A second feasibility study completed by Wright in February 1988 demonstrated that a conventional agitated leach circuit using a finer grind justified the higher capital cost of this process. This study was accepted and formed the basis of the project authorization.Preliminary engineering commenced in April 1988 and detailed engineering was started in September of that year with the objective of a mine/mill start-up by early 1990.The short time period available to meet the project deadline was additionally challenging, as site access is either by plane year round, and winter road from early February to late March. Normal lead times for equipment design, specifications, quotations and bid analyses were not possible, if the winter road time-slot was to be met."

Jan 1, 1990

By Bryan A. Bartley

"Jim Macdonald was promoted from Waterworks Engineer to Streetworks Engineer for the Wellington City Council. He thus took over the City quarry and responsibility for its performance. The crushing plant needed upgrading and he thought it would not be difficult to make a crusher to overcome a problem. If he had a longer association with the quarry industry he would have used conventional methods and equipment, so his invention arose from a fresh view of crushing.He invented the Barmac Rotopactor after observing a wear resistant chute with a lining of dead stone trapped above a piece of angle. This idea led to a rotor lined with trapped stone, rotating on a vertical spindle. In trying - a model rotor he observed the impact against a circular rill of stone in half an oil drum and thus was started a very fertile new area of developement in comminution. The writer helped in the developement and shares in the patents. Jim died in January 1982, just as the machine was becoming known and some 200 were in use in various countries. Most were working in hard rock quarries et that time but recent uses have carried the machine into the mining and mineral processing field. This paper is directed at that industry."

Jan 1, 1985

By D. Rahal, I. Shah, B. Hines, S. Reddick

"The development of stirred milling technology has proven to be an enabling technology in fine grained metallic deposits. The increased grinding efficiency allows for the economic size reduction required to produce saleable concentrates. This improvement has opened the door to retreatment projects such as reprocessing of historic gold mine tailings. Recently a VXP2500 stirred mill was installed at Casmyn Mining Zimbabwe Turk mine to allow the reprocessing of historic mine tailings. After commissioning, an extensive joint research project between Casmyn and FLSmidth Minerals was conducted to optimize the performance of the mill. This paper outlines the challenges encountered in the optimization program and the outcomes from the joint research project.INTRODUCTION The development of stirred mills has opened up many new opportunities to recover valuable minerals from refractory ores. Stirred mills have already been reported as an enabling technology for the processing of fine grained ores in lead/zinc applications (Pease, Young and Curry, 2005). Historically, conventional ball mills were unable to economically grind material fine enough to produce the required recovery or concentrate grade. This often resulted in the loss of valuable materials to the final plant tailings. The advent of stirred milling has now provided an opportunity which has both economic and environmental benefits, the reprocessing of historic mine tailings. Many tailings stockpiles now contain economic concentrations of gold because of the ability of stirred mills to efficiently and economically grind the tailings to liberation sizes. The primary advantage of reprocessing historic tailings is very low cost processing. Recovering the tailings is usually as simple as pumping the material into the reprocessing circuit. The disadvantage of reprocessing tailings is the low feed grade and greater difficulty in recovery valuable minerals at the same size distribution. The material has already been previously processed so fresh surface area or an increase in metal exposure is required. These are both provided by regrinding the tailings material."

Jan 1, 2014

By S. Feinstein, I. Reznik, Y. H. Hatzor, O. Shitrit, Y. O. Rosenberg, H. J. Vinegar, S. Nguyen

"We report the effects of porosity and total organic carbon (TOC) content on the mechanical behavior of organic-rich chalks, based on laboratory measurements. The organic-rich chalks are from the Senonian Ghareb and Mishash formations. This deposit is an immature oil-prone source rock, about 300 meters thick, situated at shallow depths (several hundreds of meters) in the Shefela basin in central Israel. The study is based on core data collected from the Zoharim borehole.Three main phases comprise the bulk volume of this rock: 1) carbonate mineralogical phase (predominantly calcite); 2) pores (mostly saturated with brine); and 3) solid organic phase (type IIS kerogen). The mechanical effects of porosity (23-45%) and TOC (4-19%) are evaluated based on bulk incompressibilities vs. volumetric fractions analysis. Volumetric fractions of mineralogical and organic phases were obtained by porosity, density and TOC measurements. Saturated rock bulk moduli were measured through ultrasonic velocity measurements. Saturated rock bulk moduli correlate best with the sum of kerogen and pore volume fractions, rather than each of them separately. The volumetric fractions were also used to define the range of bulk modulus of the organic-rich chalk, using the Hashin-Shtrikman upper and lower bounds. Based on Marion's Bounding Average Method (BAM), the normalized stiffness factor of the chalk averaged at w=0.197, indicating a soft pore geometry. Using the average normalized stiffness factor, the BAM predictions showed excellent match with the saturated rock bulk modulus and the sum of kerogen and pore volume fractions. Based on our findings, it is inferred that kerogen grains do not support the solid skeleton but are distributed between the minerals grains. We also propose here a straightforward approach for calculating the Hashin-Shtrikman bounds, based on TOC and porosity measurements. The mechanical effects of porosity and TOC indicated here are very important for the understanding of the influence of immature kerogen presence in highly porous chalks, as the Ghareb and Mishash formations."

Jan 1, 2015

By S. Haggarty

"Wholly owned by Corona Corporation, the Nickel Plate Mine is an open pit, gold mining operation located near Hedley, 50 km west of Penticton, in Southern British Columbia.Milling capacity is 3,400 SDT /Day with an average feed grade of 0.080 Oz Au/Ton. Gold processing involves three stage grinding including tertiary pebble milling, two stage cyanide leaching and filtering and Merrill-Crowe zinc precipitation.A summary of process improvements required to increase milling capacity above the initial design tonnage of 2, 700 SDT /Day are described -Installation of a 42"" x 65"" Allis Chalmers Gyratory crusher replacing a 42"" x 48"" Traylor jaw crusher.Installation of polyurethane screen panels replacing woven wire cloths on an 8' x 20 ft triple deck screen.Installation of a hopper clarifier to reduce filtrate turbidity.Automated control of solution balancing and effluent treatment circuits.INTRODUCTIONThis paper describes mining and milling practices at the Nickel Plate Mine and details post start-up process modifications which were required to attain design tonnage.Wholly owned by Corona Corporation, the Nickel Plate Mine is an open pit, gold mining operation located near Hedley, 50 km west of Penticton, in Southern British Columbia.Mining capacity from the open pit is approximately 32,000 TPD with an average waste/ore ratio of 8.4:1.Milling capacity is 3,400 SOT/Day, above initial design tonnage of 2,700 SOT/Day. Feed grade averages 0.08 Oz Au per ton yielding 85,000 troy ounces of gold per year."

Jan 1, 1991

By C. Pelley, S. Kelebek, G. Kunzel, S. Gregersen, M. Pollanen

A series of small-scale tests was conducted to evaluate the effect of micro-fracture development, imposed by blasting, on the material properties of the rock as well as its resistance to crushing and grinding. The uniaxial compressive strength, tensile strength, Poisson’s ratio, Pwave, S-wave velocity, point load test index, crushing, and grinding work indices were evaluated for samples without any blasting load and for samples which had received a variety of blast loads ranging from a single impact with a lowstrength detonating cord to a combination of impacts at different times and significant impacts with powder factors close to those used in practical blasts. It was shown that the rock mechanics parameters change even at low impacts, suggesting easier crushing of the rock. Damage values appear to be affected by timing as stress wave superposition affects the damage process. Fragmentation is achieved at higher impact levels while work index modification of the fragments is also observed when high powder factors are used, suggesting improvements in grinding.

Jan 1, 2006

By I. Spector, A. J. Smith

In 1967, BACM Industries Limited started producing walboard at Saskatchewan's first gypsum products plant, located in Saskatoon, using, as a basic raw material, gyp-sum rock supplied from the company's mine near Amaranth, Manitoba. This paper describes the geology of the Amaranth gypsum deposit, the method of mining the rock and the manufacturing facilities used to produce wall-board.

Jan 1, 1970

By James Ransford Dankwah

This work investigates the effect of waste plastics addition on the reduction of iron oxide by sawdust. Composite pellets of iron oxide containing sawdust (SD), high density polyethylene (HDPE) and blends of SD/HDPE were fired under argon in a horizontal tube furnace at temperatures between 1300 and 1500 °C. Reduced samples were weighed after 10 minutes of reduction and the extent of reduction was calculated from the measured oxygen contents of the compact. The results showed improved extent of reduction when SD was blended with HDPE. The highest extent of reduction was achieved at SD:HDPE ratio of 2:3.

Jan 1, 2014

By Adam Wright

"Les Mines Selbaie have been producing copper and zinc concentrates from the A-1 ore zone since 1986. This paper reviews major developments in flotation practice over that period. Firstly, ore blending strategies for Selbaie's non-homogeneous polymetallic ore are discussed. The evolution of reagent use is then reviewed, focussing on copper collector trials. A description of Selbaie's flotation columns follows. Finally, past and future developments for process control are outlined.Depuis 1986, Les Mines Selbaie produisent du concentré de cuivre et de zinc provenant de la zone A-1. Cette conférence présente !es développements majeurs en flottation qui ont eu lieu depuis ce temps. Premièrement, la stratégie du mélange de minerai non homogène polymétallique est présentée. L'évolution des réactifs utilises est démontrée, particulièrement !es collecteurs de cuivre. Par la suite !es colonnes de flottation sont presénteés et finalement !es développements futurs ainsi que ceux qui ont eu lieu dans le contrille de procédé sont expliqués.1.0 INTRODUCTIONLes Mines Selbaie, wholly owned by Billiton Metals Canada Inc., is currently mining and processing ore from a polymetallic surface deposit located 288 km northwest of Val D'or, Quebec (fig. 1). The mine produces copper concentrates containing gold and silver as well as zinc concentrates containing silver. 8000 dmt/day of ore is crushed and ground to 80% passing 49 µm, prior to processing by conventional copper/zinc differential flotation. The flotation circuits consist of both conventional and column cells (fig. 2). Detailed descriptions of the mill are available elsewhere in the literature (8,9)."

Jan 1, 1995

By Kazutoshi Haga, Batnasan Altansukh, Atsushi Shibayama

In this study, the dissolution behavior of copper and arsenic from five different copper concentrates containing 21-34.7 %wt. of Cu, 1.3-7.3 %wt. of As and 15.6-27.5 %wt. of Fe were investigated under high pressure oxidative leaching conditions. Leaching experiments were conducted using different leaching reagents such as ferric sulfate, sodium chloride, deionized water, sulfuric acid and ferric sulfate-sodium chloride mixture in order to compare the effectiveness of each on the dissolution of copper and arsenic from copper concentrates. The effects of ferric sulfate (0.1-1 mol/L), sodium chloride (0.1-0.5 mol/L) and leaching temperature (100-180C) were investigated under 1 MPa autoclave pressure , particle size >45 m, pulp density of 100 g/L, stirring speed of 750 rpm and leaching time of 60 min. Results showed that the most effective leaching reagents for the dissolution of copper are 0.34 M NaCl and a mixture of 0.1 M Fe2(SO4)3 - 0.34 M NaCl, in which As extraction did not exceed 7%. The copper concentration in the pregnant leach solution obtained after leaching using 0.1 M Fe2(SO4)3 - 0.34 M NaCl was 13-30 g/L, while the concentration of As and Fe were in a range between 0.3-1 g/L and 0.2-10 g/L at 160C. INTRODUCTION Copper sulfide minerals, such as chalcopyrite (CuFeS2), chalcocite (Cu2S) and covellite (CuS) are used for the production of copper in mining and metallurgical industries worldwide. The conventional pyrometallurgical processing followed by electrorefining is a common process for the production of high purity copper (Moskalyk, & Alfantazi, 2003; Padilla, Aracena, & Ruiz, 2012). However, the degradation of copper content and an increase in the amount of metal impurities, especially arsenic (As), antimony (Sb) and bismuth (Bi) in copper concentrates cause serious problems to the production of copper and environmental pollution (Basha, Selvi, Ramasamy, & Chellammal, 2008; Long, Peng, & Bradshaw, 2012). Hydrometallurgical processes are used to recover metals from refractory ores, concentrates and tailings by leaching (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018). During the last decades, hydrometallurgical processing for the recovery of copper from enargite-bearing copper ores and concentrates has been extensively investigated due to its relatively more environmentally benign and less capital costs than the conventional pyrometallurgical processes (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018; Choubey, Lee, Kim, & Kim 2018). However, several issues facing the hydrometallurgical processing are incomplete extraction of copper associated with the coprecipitation of iron compounds, difficult stabilization of toxic impurities and production of a contaminated copper product (Dreisinger, 2003; Safarzadeh, Moats, & Miller, 2014). Hence, a number of studies have been conducted on the selective dissolution of copper from copper ores and concentrates or selective removal of arsenic leaving a clean copper residue using various leaching reagents or lixiviants (Safarzadeh, Moats, & Miller, 2014; Mihajlovic, Strbac, Zivkovic, Kovacevic, & Stehernic, 2007; Tongamp, Takasaki, & Shibayama, 2009). More extensive studies on the selective extraction of copper from enargite- bearing copper ores and concentrates are much needed to improve copper recovery.

Jan 1, 2019

By P. J. Morgan, E. Wolmarans

The selection of the main mineralised zone (MMZ) concentrator milling circuit has followed a comprehensive route over the preceding 10 years. Many knowledgeable parties have contributed and assisted in developing the milling circuit according to the best available information. The final selection of the flow sheet was completed after DRA conducted a techno-economic study on four possible comminution circuits ranging from conventional crushing and ball milling to fully autogenous (AG) milling. The results from all the above mentioned work indicated that it is possible to treat the MMZ ore at the desired throughputs to the desired grind sizes in a fully autogenous grinding circuit. The flow sheet was designed to treat the main mineralised zone (MMZ) arising’s from both underground and open cast operations and incorporates a fully autogenous grinding (AG) circuit. Commissioning of the 375ktpm concentrator at Nkomati Nickel began in August 2009 with the first ore reporting to the mill on the 15th of September 2009.

Jan 1, 2012

By Vincent Roy

Numerous on going activities at the Daniel-Johnson Dam : ? Surveillance / monitoring ; ? Structural investigations & evaluation ; ? Injection work ; ? Crack mapping ; ? Instrumentation management (drawings & inventory) ; ? Drainage system state ; ? Foundation evaluation

Aug 1, 2013

By Sebastián H. Troncoso

Open pit mine planning is normally based on the selection of an economic envelope that has been generated by the Lerchs & Grossman algorithm, which requires individual economic valuation of each block of a geological block model. The selected envelope, usually called Final Pit, is later discretized in pushbacks and benches, which correspond finally to the fractions that the production schedule will allocate in time. This planning methodology has been widely used by the mining industry in countless projects to exploit open-pit mines, however, in many of the deposits can be found some undesirable elements such as arsenic, phosphates, ash among others that tend to reduce the value of elements of economic interest, either by disrupting the processes of concentration or as deemed impurities that affect the quality of the final product. Thus, a fair amount of blending would be required to capture the real value of the ore body. The traditional approach under estimates a block value by penalizing the block value by the amount of contaminant even when at the mine a set of blocks is amine at a time. This paper proposes a methodology for computing cut off grades over time that would maximize net present value allowing blending that should fulfill a set of metallurgical constraints. Preliminary results show significant increases in value and reserves quantity, questioning the use of fixed cut-off grades in generating a production schedule when blending applies to fulfill a set of geometallurgical constraints. Finally, the model allows the authors to show the importance of having adequate planning tools that will lead to the ability to look at the potential economic value of a mineral deposit.

Oct 1, 2009

By J S. Kermeen

This paper deals with drilling and blasting techniques at the Phoenix mine, near Grand F1nks, B.C., which is now operated by The Granby Mining Company Ltd. The difficulties of open-pit mining through former underground workings are unique to this operation. Rock types are related to drilling conditions, and the methods used, or contemplated, to overcome the several difficulties are outlined. The equipment used is also mentioned. Finally, blasting techniques and costs are reviewed. Introduction T HE Granby Mining Company operates on open-pit copper mine and concentrator at the site of the original Phoenix mine, near Grand Forks in south-central British Columbia. The pit is a side-hill cut measuring approximately 1,000 by 2,000 ft., and covering a vertical range of from 4,550 to 4,950 feet above sea level. The mill is currently operating at 1,900 tons per day. A 4.5 to 1.0 waste to ore ratio necessitates the removal of some 15,000 tons of ore and waste per working day. Mining conditions at Phoenix are extraordinary in that all the ore zones currently being

Jan 1, 1965

By S. McPherson, M. Chalkley, R. Ellenwood, W. Nicot

"The Moa JV was formed in December 1994 as a 50/50 joint venture between Sherritt Inc. and General Nickel Company S.A. of Cuba. The joint venture consists of the mine and Pedro Sotto Alba processing plant in Moa, Cuba and the Corefco refinery in Fort Saskatchewan, Alberta, Canada. Throughout its now nearly 23-year history, the Moa Joint Venture has faced a number of global, operational and process challenges. Global challenges have included high input costs such as for fuel oil and sulphur, as well as low nickel prices. Operational and process challenges have included increasing ore haulage distances, decreasing ore Ni + Co grades, and increasing ore impurity contents. For example, increases in impurities such as Mg and SiO2 in the ore result in increasing acid demand in leaching and decreasing solids content of preheated leach feed slurry, respectively. In spite of these challenges, the mixed sulphide production capacity from the Moa Joint Venture has increased from 12,500 to 37,000 t/y Ni + Co since 1994 by judiciously targeting expansion projects and by continuously performing process and operational improvements.INTRODUCTION The Pedro Sotto Alba high pressure acid leach (HPAL) plant at Moa Bay, Cuba has a long history (Maslenitzkiy, 1969; Chalkley, 1996; Carlson, 1960; Roy, 1961; Carlson 1961). The plant was constructed by Moa Bay Mining Company and started up in late 1959 after several years of process development by Freeport Nickel Company, including operation of a 10 t/d plant in Hoskins Mound, Texas and a 50 t/d plant in Port Nickel, Louisiana (Southwick, 2004), as well as development work by Chemical Construction Company (Chemico) to precipitate metal sulphides (Roy, 1955a; Roy, 1955b). The Moa plant had an original reserve of 50,000,000 t of ore, graded at 1.35 wt% Ni, 0.146 wt% Co, 3.70 wt% Si, 1.00 wt% Mg and 4.5 wt% Al. At original design rates of 2,000 kt/y of ore, the design capacity of the plant was around 25,000 t/y Ni + Co as a mixed sulphide product, which would be shipped to the Port Nickel Refinery in Louisiana for processing. In 1960, the plant changed ownership and was restarted in 1961 under Cuban management. Between 1961 and 1989, the plant was also supported by involvement from the former USSR, supplying spare parts and consumables. In 1994, Sherritt Inc. partnered with the General Nickel Co. S.A. of Cuba and formed the Moa Joint Venture. This joint venture includes the mine and Pedro Sotto Alba HPAL processing plant in Cuba and the Corefco refinery in Fort Saskatchewan, Alberta, Canada. Under this partnership, ore is processed in Cuba to produce a nominally 55 wt% Ni and 5 wt% Co mixed sulphide product that is shipped to Canada and processed by Corefco to produce class 1 nickel, high grade cobalt and byproduct ammonium sulphate fertilizer."

Jan 1, 2017

By R. Simkus, J. K. Gowans

Polaris Operations, a Cominco Ltd. lead-zinc mine, is located on Little Cornwallis Island, Northwest Territories, at Latitude 75°23' N and Longitude 96°75' w. When the Polaris Mine went into operation in late 1981, it became the world's most northerly base metal mine.The Polaris Concentrator, housed in a unique process barge, is a basic lead-zinc concentrator where flotation is used to produce the lead and zinc concentrates.The paper describes the ""coarse lead"" flotation feature incorporated in the concentrator flowsheet and describes the start-up problems and continuing operation of this circuit.

Jan 1, 1984

By Karl Nienhaus

The BGMR has developed a method for an online-analysis during drilling processes in cooperation with Fraunhofer Institute for Laser Technology in Aachen and various industrial partners. This allows for efficient utilization of deposits, maximum recovery of desired minerals and frequent and costefficient updates of deposit models. The used method, namely laser-induced breakdown spectroscopy (LIBS), was implemented into a compact demonstrator in order to analyze drill dust and loosened material within mining environments and directly on drill rigs. The consequently developed OFURanalyzer (OFUR ? online-analysis of mineral deposits) provides possibilities for exploration and visualization of deposit models in real-time and thus creates an opportunity for innovative applications, such as contactless, continuous quality control of minerals throughout the extraction process. A major challenge in the execution of this project was the necessity to develop a laser system that could withstand rough mining environments dominated by dust, shocks, vibration and therefore would be suited to operate under such conditions.

Oct 1, 2009