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By Daniel Hodouin, Bruno Girard, Claude Bazin

"Mineral processing and smelting operations have the reputation of being complex processes. Process operators of these unit operations have acquired from experience and basic training a superficial understanding of the relationships between process output and manipulated variables. This understanding, which is a very valuable know-how for process control and optimization as well as for operator training, is however difficult to extract. This paper describes a time series based approach that can be used to extract this know-how, which can subsequently be used for operator training. The method is illustrated with an example taken from a smelting process.IntroductionThe emergence of Total Quality Management and Statistical Process Control philosophies in the North American industry (Brown, 1991 ), as well as the increasing popularity of the IS0- 900x standard accreditation, provide a strong incentive to review the training procedures for plant operators. The importance of employees training is acknowledged in two of the 14 points of Deming's road to quality (Deming, 1982) as well as by control engineering consultants (Hendricks et al., 1994). However guidelines for the operation of industrial processes are often transferred from an operator to another one with allowance for flexible interpretation of many operating rules. Such transfer is typical of complex industrial processes to which particulate processes of ore dressing and extractive metallurgy belong.Capturing the knowledge of experienced operators is often a difficult task. Cognition engineers are able to extract this knowledge and transfer it into rules that ultimately are implemented into expert systems. Cognition engineers are however not always available and plant supervisors may prefer to keep training as an internal liability. Another difficulty through interviewing of the process operator is its lack of willingness to transfer what he has learned by hard working to other people, whose ultimate goals are to develop computerized applications that may replace him in a near future, or reduce its importance in the production line.This paper shows that time series can be advantageously used to capture the average process understanding of plant operators and to help reviewing training procedures. The method s illustrated with data collected during normal operation of electric arc furnaces and it is shown how the approach has led to a fruitful discussion between process engineers and operators that ultimately provided savings of more that $200,000 per year per furnace. There are 9 furnaces in the studied foundry. Such savings are not the results of application of advanced control strategies nor of ""miracle press button software», but of an improve awareness by engineers and operators of the importance of properly defining the operating procedures to control a process."

Jan 1, 1996

By A. Nielsen

"The Polaris mining operation (Fig. I) is located on Little Cornwallis Island in the Canadian Arctic, an area with a very limited shipping season and a hostile climate. The operation consists of the underground works, barge-mounted concentrator facilities, concentrate storage and shipping facilities, a personnel accommodation complex and an airstrip.The electrical energy for the operation is provided by four 2.3 MW diesel-generators installed on the barge. The power is distributed through a 4.16 KV system connected directly to the generators and to the large motors. All other loads are supplied at 600 V or 120 V.The power plant has been designed as a Total Energy System with maximum energy efficiency and thus designed for maximum heat recovery. The waste heat recovered on a year 'round basis is expected to be sufficient for the process heating and drying as well as the space heating for offices and accommodations. The engine cooling system consists of two glycol-water loops, a "" cold"" loop for charge air and lube oil cooling and a ""warm"" loop for jacket water cooling. Each loop is complete with heat exchangers f or heat recovery and out-door forced air radiators to dispose of excess heat. A water loop in the concentrator area transfers the heat from the jacket water loop and the exhaust gas boiler to the space heating equipment, the process and the heating system for the accommodations. Part of the exhaust gas is diverted to the concentrate dryers.IntroductionThe Polaris Mine is a highly mechanized, trackless mine with access through a decline. Drill equipment is electrically powered to minimize fumes and heat. Dry development and production drilling eliminates the need for an underground water supply system.The concentrator, power generators, work shops and auxiliary facilities are located on a barge which is beached in a prepared berth on the shore of Crozier Strait. Oil storage tanks, concentrate storage and accommodation buildings are built on shore. The concentrate storage building ha s a 12-month capacity and is fed from the concentrator by a conveyor system.The accommodation building houses up to 240 people, includes administration office, dining and recreation facilities. The complex is housed in on e interconnected building with public use areas alternating with accommodation units.At full production annual concentrate shipments will amount to about 220,000 wet tonnes. This will be accommodated by about 8 shipments during the regular shipping period-August 15 to October 15. Re-supply of general cargo and fuel will be carried out annually in about 3 shipments within the same period."

Jan 1, 1984

By Margaret Armstrong, Normand Champigny

"In order to define the ore/waste boundaries as finely as possible, mine planners often insist on kriging very small blocks (relative to the spacing of the available sample grades). These kriged estimates for very small blocks are over-smoothed. For several simple sampling patterns the relation between the block size and the variogram range, and the kriging results were studied. Meaningful estimates of individual blocks are obtained when the variogram range is large compared to the block size and the sample spacing. For a variogram range less than half the sample spacing, the kriged block estimates were found to be uncorrelated with the actual grades. The over-smoothed estimates should not be used for calculating recoverable reserves IntroductionNowadays most mining companies know that the geostatistical estimation method, kriging, gives the best estimates of individual block grades. As most mining engineers and geologists are now familiar with the basics of geostatistics, the definitions of the variogram and of kriging will not be repeated here. An appendix containing definitions and mathematical formulae is given at the end of the article for the sake of completeness.Even though it is well known that the kriged estimates for very small blocks are over-smoothed, mine planners who want to define the ore/waste limits as finely as possible (especially for selective underground mines), are often tempted to krige very small blocks. So we ask how small is ""too small""? Clearly ""smallness"" cannot be expressed in yards or metres. It depends on the nature of the orebody under study and this is reflected in the variogram model chosen to describe it.To answer this question, the changes in four key parameters are studied:• the kriging variance,• the variance of the estimated grade,• the slope of the linear regression of the true but unknown grade on its estimated value, and• the coefficient of correlation between the actual grade and the estimate;as a function of the block size and the variogram range. (Definitions of these four parameters are given in the appendix). Ideally the kriging variance should be low while the variance of the estimated grade should be equal to the dispersion variance of the blocks. The slope should be l.0, and so should the correlation coefficient.These four parameters have been calculated for several simple 2-D examples. The effect of using kriged block estimates that are too heavily smoothed are shown, and the dangers of trying to use these estimates for predicting recoverable reserves for small blocks (i.e. selective mining units) are pointed out."

Jan 1, 1989

By Q. J. Xu

In blasting sensitive areas, the use of traditional blasting for rock fracture is restricted. Soundless cracking technology has been presented to achieve rock fracture instead of blasting. However, when the diameter of a borehole is larger than 50 mm, it is very dangerous that the soundless cracking agent (SCA) will have a tendency to gush from the borehole, which also leads to cracking failure. In this paper, a new large scale borehole soundless cracking technology is presented, using a new enlarged drill bit and mechanical plugging, which could solve the hole gush problem. Experiments were made to measure the expansive pressure of the SCA working in boreholes of F40 mm and F100 mm. The difference between two sizes of cracking phenomenon was compared and analyzed to show that large scale borehole soundless cracking has larger expansion pressure and faster reaction speed than traditional soundless cracking. Based on the experimental data, finite element analysis was utilized to study the stress concentration and stress distribution between two neighboring large boreholes in several media during the cracking process. Using the data from experiments with finite element numerical simulation, the brittle material fracture condition was analyzed to design the parameters of the borehole mesh. With the theoretical analysis, the technological process and reference parameters of large scale borehole soundless cracking were proposed, including the borehole spacing, burden distance, depth of the borehole and so on. Then a single rock was cracked with the new enlarged drill bit and new plugging, and the results proved the feasibility of the new large scale borehole soundless cracking technology, which achieved the expected goal of improving the cracking ability and reducing the cracking time. This paper presents a new efficient rock cracking method for blasting sensitive areas, and which could apply to mining engineering, demolition engineering, excavation engineering and other engineering applications where there is a demand soundless and non-vibration fracturing of rock.

Aug 1, 2013

By J. A. Haskin

The Engineering Department at Flin Flon has the following responsibilities: 1. Underground Surveying. 2. Open-Pit Surveying and Supervision of Blasting. 3. Mine Geology and Examination of Prospects. 4. Design of Plant Additions and Improvements. 5. Miscellaneous Field W ork, Reports, etc. The entire surface plant, as well as the underground workings, are located by co-ordinates based on a triangulation system covering an area approximately one mile long by one-half mile wide. The origin of this system and one end of the base-line is the southwest corner of Township 67, Range 29, west of the Principal Meridian. The bearing of the base-line was determined from the township boundary, which runs true north, so that all other bearings are almost astronomical. This is a great convenience when our surveys must be tied to mineral claims, rights-of-way, and land subdivisions. As the base-line extends across Flin Flon lake, its measurement across smooth ice was simple and accurate. The length, slightly over 1,600 feet, has been adjusted for the temperature and tension of the standard tapes used and checked sufficiently to ensure a maximum error of 1 : 15,000. Each angle was repeated ten times, but as no special refinements were used, errors of as much as four or five seconds are possible. The elevation of the origin of the triangulation system was approximately 1,050 feet above sea level. The use of this true elevation in our surveys would have resulted in negative elevations underground as soon as the mine reached a depth greater than 1,050 feet. To avoid this complication we therefore added 4,000 feet to this elevation. The collar of the Main shaft, for example, which is actually 1,072 feet above sea level, appears in our records as at elevation 5,072. Underground surveying, which follows standard practice, is done by two mine surveyors who share three helpers. The meridian is taken underground from two wires hung in either of the two vertical shafts.

Jan 1, 1935

By S. E. Malouf, W. Thorpe

"The Chibougamau Explorers property comprises 35 claims or 1,726 acres in La Dauversiere and Rohault Townships along the west shore of Norhart Lake. The property is serviced by a twelve-mile all-weather high-way leaving the main St. Felicien-Chibougamau highway at mileage 121. This road junction will be a railway siding on the C.N.R. St. Felicien-Chibougamau branch line in early 1957. The property lies 30 miles due south from the town of Chibougamau. The mine-sponsored Chibougamau Explorers town has a population of 250 with 28 residences and a well equipped recreation centre.The original showing was found by Rider Storn, a prospector financed by Val d'Or interests during the 1950 field season. This showing did not stand up to detailed prospecting but it did lead to the discovery of the main zone by A. Oakley during the following winter.The company was incorporated in I 950 under the presidency of E. Ayrhart. Financing following the initial drilling involved an increase of capitalization from 3 to 5 million shares and the advent of financing and management control by Anacon Lead Mines Limited.Exploration and development to Jan. I, 1956, has included 42,331 feet of surface diamond drilling; 31,363 feet of underground drilling; a 3-compartment shaft to 1,245 feet, with 8 levels at 150-foot intervals; and I 1,054 feet of drifts, crosscuts, and raises. Ore reserves to the 900-foot level, proven and partly developed, total 546,725 tons averaging 0·30 ounces gold and 0·93 per cent copper. A mill with capacity of 500 to 750 tons will be installed by January 1956.GeologyThe Chibougamau Explorers property lies along the belt of Keewatin volcanics striking through Bachelor Lake, Opawica Lake, Doda Lake, Coapatina, Norhart Lake, and Rohault Lake, where it is in close proximity to the paragneisses of the Grenville series. The boundary of the paragneisses, a gradational contact, lies within a mile of the Chibougamau Explorers shaft collar."

Jan 1, 1949

By J. R. Pronovost, P. M. Tiley, D. W. Watts

"In 1980, J.S. Redpath Limited were retained by Dickenson Mines Limited to increase hoisting capacity from 700 to 1000 tons per day. The hoisting capacity of the present skip-over-cage system could not be increased significantly through design modifications. The other constraint was a minimum of disruption to the current production. The Dickenson shaft was a three-compartment shaft from surface to the 8th level and four compartments before the 8th to the 23rd level. The solution to the problem was to remove the services from the shaft man-way compartment from the 8th level to surf ace by drilling a borehole close to the shaft and rerouting those services through the borehole to the 8th level. A service compartment would then be available from surf ace to the 23rd level.By providing a service hoist in this compartment, the skip hoist would then be free to produce the 1000-tons/day required production. This paper will describe the search for the required hoist and the friction hoist selected for the project, and trace the design modifications required to modify the hoist to suit the Dickenson application and satisfy the Department of Labour safety requirements.The paper has been co-authored by J.S. Redpath Limited and C.L. Tiley Associates, who participated in the design changes and eventual commissioning of the hoist.IntroductionIn the earl y part of 1980, J.S. Redpath Limited conducted an engineering study on the Dickenson Mines Limited No. 1 Shaft materials handling system. The object of the engineering study was to investigate the possibility of increasing the tonnage hoisted from the No. 1 Shaft complex by adding a service cage compartment and service hoist. This would permit utilization of the existing hoisting plant solely for production hoi sting. A further consideration of the engineering study was that any alterations had to be accomplished with little, or no, interruption to production."

Jan 1, 1982

Adaptive Processing as proposed in geometallurgy has to rely on spatially interpolated information on geometallurgical parameters like phase composition, size distributions of particles of different phases, grain shape parameters, and portions of value elements in different grains. Using the geostatistically predicted values for adaptive processing, e.g. for the selection of milling diameters, thresholds in physical separation, or choices on using an extra pre-separation step, is typically not optimal. Mathematically this effect is introduced by two forms of nonlinearities: 1) The nonlinear scales of compositions, distributions, and shapes have special properties with respect to geostatistics. Classical geostatistics creates some artefacts for these nonlinear scales. On the other hand, modern geostatistical procedures adapted to these scales do not provide unbiased results with respect to linear transformations of the data (e.g. biased block estimates). 2) Neither economic nor ecological effects (e.g. monetary gain) of processing decisions are linear in the interpolated geometallurgical parameters. These nonlinear transforms are not unbiasedly estimated by the likewise transformed unbiased geostatistical predictions of the geometallurgical parameters. Furthermore, we need to optimize the conditional expectation of the gain, rather than obtain an unbiased estimate. Standard geostatistics as such does not provide the "right sort" of estimates for adaptive processing. A nonlinear kriging procedure is needed to approximate the nonlinearities mentioned before. We propose to solve these problems simultaneously using a nonlinear geostatistical technique for predicting the target function (the monetary gain), rather than to predict the geometallurgical parameters and compute the gain from them. The optimization can then be performed directly on this estimated function. It can be shown that this optimization performed on the conditional expectations, not on unbiased predictions, would yield the best possible processing choice. We propose a procedure choosing the processing parameters on an approximation of the conditional expectation. The difficulties with the classical approach and the effectiveness of this new approach are illustrated by a simplified simulation example with a single processing parameter and a simple dependence on the microstructure.

Aug 1, 2013

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 Steven R. McCutcheon

"The Bathurst-Newcastle district, a subcircular area approximately 50 km in diameter, contains 37 massive-sulfide Zn-Pb-Cu deposits that have defined tonnages, and has 58 additional significant occurrences. The largest deposit is Brunswick No. 12, with total reserves (including past production) of 148 million tonnes grading about 12.5% combined Pb-Zn. The deposits/ occurrences are spatially associated with a felsic volcanic pile that contains a significant sedimentary component. In the vicinity of the Brunswick deposits the pile is divisible into lower and upper sequences. The lower is characterized by megacrystic rhyolite (QFAS) and coarse- to fine-grained quartz (± feldspar) wackes and/or tuffites (QAS); the upper sequence is characterized by aphyric to sparsely porphyritic rhyolites (HWR), micro-breccias, and acid hyaloclastites with interbedded dark gray sedimentary rocks. A greenish gray to dark gray siltstone (GGS) separates the two sequences and hosts the sulfide deposits. Lithogeo chemistry of least altered samples shows that the QFAS and HWR have an igneous character, whereas the QAS and GGS have a sedimentary signature. The two types can be distinguished on the basis of total alkalies (Na2 + K2O) versus SiO2/Al2O3 and Ti02 versus Fe/(Fe +Mg). Some trace elements, such as Sr and Y, also help differentiate the two types. The QFAS differs from HWR by having higher La/Yb ratios but lower Zr and Th. The Bathurst deposits do not show a close spatial association with rhyolites but rather are stratiform bodies Oacking clastic textures) within fine-grained sedimentary rocks. The scarcity of barite and other sulfates indicates that there was little mixing between the metal bearing fluids and cold seawater. High Fe/(Fe+Mg) ratios in chlorites also indicate the fluids did not contain a significant seawater component. These features and the large size of many deposits suggest that the deposits formed from metal-rich pore waters that were released catastrophicallyfrom an acquifer in the footwall stratigraphy."

Jan 1, 1992

By T. Gonzales, M. Maleki, M. Caçador, B. Climenhaga, H. Ghorbani

The campaign life of furnaces in the copper making industry is greatly affected by damage caused by normal operation cycling and abnormal conditions. Proper monitoring and assessment of the existing conditions and safe remaining life of the vessels and their refractory lining is key to prevent an unplanned early replacement. This paper presents an industry-proven refractory management program as well as a typical fitness-for-service assessment of a rotary vessel shell in order to extend the campaign life of copper making furnaces and enhance safety. The furnace refractory management system involves a comprehensive audit and monitoring program which has been successfully applied to several copper flash smelting furnaces over the past 15 years. The fitness for service assessment provides a methodology to estimate the remaining life of the steel components by evaluating the existing damage and future damage rate based on the historical survey data. The above practices can provide the operator with cost savings by extending the vessel campaign life in lieu of implementing more rigorous refractory lining and shell audits. INTRODUCTION Primary copper smelting is generally divided into two categories of flash smelting and open bath smelting. There are many different types of vessels that can be used to convert the copper concentrate to blister copper including, flash, Mitsubishi process, TSL (ISA/Ausmelt), SKS Bottom Blowing, Anode and Teniente/Noranda furnaces, reactors, and Peirce Smith converters. Each vessel has its own complex and intense environment which requires careful refractory and steel design to withstand the severe thermomechanical and thermochemical attacks within its campaign life. The campaign life of furnaces is greatly affected by damage caused by normal operation cycling and abnormal conditions. Proper assessment of the existing conditions and safe remaining life of the vessels and/or refractory lining is key to prevent an unplanned early replacement. This paper presents an industry-proven refractory management program as well as a fitness-for- service assessment methodology to estimate the remaining campaign life of copper making furnaces resulting in cost savings and enhanced safety of the operating vessels.

Jan 1, 2019

By M. R. Henzinger, W. Schubert

"The quality of the backfill applied at a shield driven tunnel is of significant importance failure of the lining. The used material also affects the interplay between support and excavation behaviour.In this research the state of the actual bedding situation is investigated by an analytical approach, laboratory tests and numerical simulations. The numerical simulation using continuum methods have shown shortcomings. These shortcomings regarding the instability of large strain simulations restrict the possibility to reproduce the relocation of the backfilling material. The performed scaled model tests have shown that the regripping process of a double shielded tunnel boring machine affects the soil failure within the gap. With the analytical results the influence of the wall friction can be predicted. Therefore the angle of the failure surface can be determined. The results have shown that the wall friction significantly increases the failure angle. INTRODUCTIONFor tunnel boring machines (TBM) with single (SM) or double (DSM) shield, precast liningsegments are used as rock support. These segmental linings are installed within the protection of the TBM shield. In order to reduce the friction while advancing, the shield diameter is smaller than the diameter of the cutter head and therefore also smaller than the excavation boundary. The difference depends on the expected radial convergences. Since the lining segments are mounted within the shield the steering gap between the shield and the excavation boundary increases and is hereinafter called annular gap.In the process of tunneling in hard rock, the annular gap is usually filled with fine grained and closely graded gravel, termed as pea gravel. The backfilling process starts as soon as possible after the shield tail passes the lining. Due to the operational procedure within the working area of a TBM, a fully backfilled annular gap might not be established after every ring closure. This leads to an unfavorable distribution of the pea gravel leaving the segmental lining partially without bedding. Segmental linings can develop sufficient resistance only when properly bedded."

Jan 1, 2015

By Robert E. McIvor

"Use of the Bond Work Index to size industrial grinding equipment was widely known. Bond also described its application for measuring grinding circuit efficiency. Bond Work Index efficiency also has been used extensively by industry, but until now without concise standard methods, or a formal guideline, for doing so. The Bond Efficiency Sub-Committee of the Industrial Comminution Efficiency work group of GMSG has produced such a guideline, which is appended hereto. It covers the methods and provides examples of how to calculate the Bond Work Index Efficiency of most industrial grinding circuits. Bond Work Index testing equipment and methods are also provided.Comminution is responsible for a major portion of mineral processing plant energy consumption and operating costs. Bond Work Index efficiency provides valuable means, first, for benchmarking one’s plant performance against others in the broad industry, and hence defining the extent of the opportunity for improvement. Secondly, improvements in circuit efficiency with implemented changes can be quantified and monitored as part of continuous improvement efforts. These two important business activities are thus facilitated, and, in the absence of any other accepted method, otherwise would be lacking without the existence and application of this standard for comparison. While often loosely applied in the past by plant metallurgical staffs, the increased precision offered by this guideline provides much clearer definition of both the conceptual and statistical meaning of these measurements. Even a small improvement can be enormously financially important, but only if differentiated by measurement. This guideline was thus intended to provide industry improved means to differentiate comminution circuit efficiencies. Carbon footprint reduction is coincidental, but nevertheless can be substantial.The next steps are: (1) to establish standard reference testing materials and one or more laboratories to serve as bases for reference testing; (2) to establish the accuracy of relative circuit Work Index Efficiency determinations; and (3), to develop and provide a public data base of industrial comminution circuit Bond Work Index Efficiencies."

Jan 1, 2016

By J. A. Eisele, D. L. Pool

"The Bureau of Mines has developed a particle agglomeration technique to improve the flow of leaching solution through heaps of low-grade gold-silver ores which contain clays. Benchscale experiments and industrial practice have shown that the percolation rate of cyanide leaching solution is greatly enhanced by mixing the ore with Portland cement as a binder, moistening the mixture, mechanically tumbling to form agglomerates, and aging before heap building and leaching. A slight modification ofthe technique allows its application to precious-metal-bearing tailings, which cannot normally be treated by heap leaching because they are too fine. IntroductionMany low-grade gold and silver deposits and tailing materials from former mining operations in the western United States are potential sources for precious metals. These sources are too low grade or too small to support the capital cost of a conventional cyanidation plant but have stimulated the development of lowcost procedures, such as extraction of the precious metals by heap leaching with cyanide. The Bureau of Mines has investigated methods to improve the recovery ofgold and silver by heap leaching these low-grade materials.For heap leaching to be successful, the crushed and stacked ore must have good permeability so that uniform distribution of the cyanide leaching solution through the heaps can be achieved. Some of the most difficult gold and silver ores to treat by heap leaching are those containing excessive amounts of clays, or fines generated during crushing. Excessive amounts of slimes, generally minus 50-micrometre or 270-mesh Tyler sieve size particles, in the heap feed will slow the percolation of the leaching solution, cause channelling, or produce dormant or unleached areas within the heap, resulting in long leaching periods and low extractions. In extreme cases, the clays or slimes can seal the heap and cause the leaching solution to run off the sides of the heap rather than through it."

Jan 1, 1987

By Wei Zhang, Jinhong Zhang

"The adsorption of ethyl xanthate on copper sulfides, i.e., chalcopyrite, chalcocite and bornite, has been studied in situ in aqueous solutions by applying an atomic force microscopy (AFM). AFM images show that xanthate, such as potassium ethyl xanthate (KEX), adsorbs strongly on chalcocite and bornite mainly in the form of insoluble metal xanthate, which binds strongly with the mineral surface without being removed by flushing with ethanol alcohol. This adsorption mechanism is totally different from the one obtained with chalcopyrite/xanthate system, during which oily dixanthogen is the main adsorption product on chalcopyrite surface. The findings obtained from the present AFM study suggest that an AFM image analysis is a powerful tool for the study of the collector’s adsorption on mineral surface and as well the development of novel collector for copper sulfide flotation.INTRODUCTIONIn mining industry, copper sulfide minerals, i.e., chalcopyrite, chalcocite and bornite, are usually separated selectively from non-valuable gangue by froth flotation, which treats the highest throughput and produces the maximum economic outcome of any surface chemistry process. In flotation, collectors are added into the pulp (mixture of water and fine-grained minerals) to selectively adsorb onto the target mineral and render its surface with high hydrophobicity, which is generally beneficial for a strong mineralbubble attachment and thus a high flotation recovery.The adsorption of collector on mineral surface, therefore, is vital for a successful flotation process and it has been studied by applying various surface analysis techniques, such as IR (Infrared Spectroscopy), CV (Cyclic Voltammetry) and XPS (X-ray Photoelectron Spectroscopy). (Poling, 1963; Allison et al., 1972; Mielczarski et al., 1984; Leppinen et al., 1989; Kartio et al., 1996) These studies have revealed a lot of information, such as the reaction, product and mechanism, of the adsorption of chemicals on mineral surface. It is also of great interest to directly get the image of adsorbed collectors on mineral surface changing with pulp chemistry, such as pH, redox potential, ionic strength and chemical’s dosage."

Jan 1, 2016

By V. Nudo, R. del Villar, A. R. Laplante

"Fine iron ore recovery from the tails of Les Mines Quebec Cartier was investigated using Reichert trays and spirals. The effect of feed rate, density and splitter setting on the performance of rougher, cleaner and recleaner stages of a 2DS tray assembly was determined using 2- and 3- factorial experimental designs. The tray concentrate, assaying 27% Fe, was further upgraded by spiral to yield a concentrate assaying 55% Fe. Further laboratory scale testing showed this material to be easily upgraded to 65-66% Fe with high Fe recoveries.This paper presents the metallurgical results of the three tray stages. The model used to predict the performance of a full recovery circuit is then described. Simulation results are presented and their reliability assessed. Finally, an economic evaluation of the proposed scavenging circuit is summarized.INTRODUCTIONThe Mount Wright plant (UWP), located near Fermont, Quebec, processes daily 130,000-150,000 tonnes of iron ore assaying between 27 and 33% Fe, to upgrade it with three stages of 44.5 cm pitch GEC spirals to 66% Fe, while rejecting approximately 80,000 tonnes assaying between 6 and 11% Fe (1-3). Heavy liquid separation (at 3. 3 g/mL) of the plus 37 µm fraction (+400 mesh) of the tails reveals that a significant fraction of the iron is in the sink product, below 150 pm (100 mesh), as shown in Table 1. Also shown is the iron grade of the sink and float fractions. The high grade of the fine sink indicates that hematite ls well liberated below 105 µm (150 mesh), as previous studies have shown (4,5). The iron grade of the floats is very low in the fine size range, but increases with increasing size above 105 pm, because of incomplete liberation. Overall, two thirds of the plus 37 µm iron ls in the sink product, which constitutes less than 10% of the total tail."

Jan 1, 1987

By Brian McIntrye, Bo Trygg

"The need for grinding mills of ever increasing capacity and greater feed openings has tended toward mill trunnion diameters being beyond the size limits for roller bearings.Hydrodynamic plain bearings need to be wide enough to achieve the necessary bearing pressure in order to obtain an oil film with sufficient carrying capacity at low rotational speeds. This greater width increases edge loading resulting from the deflection of the mill under load.Because of the high surf ace pressure involved, the life of support rollers and girth rings is often inadequate-fatigue leads to flaking on the contact surfaces. Heat treatment of the girth rings also has its problems.The hydrostatic shoe bearing permits unlimited trunnion diameter and a high carrying capacity. The self-aligning shoes compensate automatically for radial deviations and large deformations can be tolerated at the girth ring.The sliding surfaces are completely separated by oil under pressure, enabling negligible starting torque and good damping capacity to be achieved together with low rotational speeds.IntroductionHydrostatic bearings have been known within machine technology since 1865, when the Frenchman Girard received a patent for such a bearing. During the following 75 years, these types of bearings were used sporadically for thrust bearings, primarily in water turbines. The first usable radial bearing was not designed until 1938. Since the 1950s, when hydrostatic bearings started to come into use within the field of machine tools, extensive research and development work has been carried out at various institutions and research centres throughout the world. Many support problems are now solved by means of hydrostatic plain bearings and the primary factor in deciding to use them is their unique characteristics:• ability to carry high loadings• no starting torque• large damping capacity• no wear• large running accuracyThe hydrostatic shoe bearing has been developed specially for use in the heavy machine industry, with the requirements for high loadings and large support diameters which this market demands."

Jan 1, 1982

By Leonard A. Wolff

"The 1980sfor the mining industry can be characterized as the survival of the fittest. High interest rates, low commodity prices, foreign subsidized competition and decreasing ore grades have contributed to a streamlining of the domestic mining industry in the United States, leading to the shutdown of operations or large cutback of work forces. The purpose of this paper is to look at the impact of a layoff on the employees of what was once a major uranium producer in south central Wyoming, twenty months after the final layoff and mine shutdown. INTRODUCTIONDecember 19, 1985, was the date of the final of three successive layoffs over a two-month period, leading to the complete shutdown of Petrotomics Company, a subsidiary of Getty Minerals Company. The final shutdown was due to a combination of depressed U3O8 prices and the takeover of Getty Oil Company by Texaco in January 1984.The layoffs affected 104 employees and included mine personnel, mill personnel and staff. Both hourly and salary workers were involved. The operation was considered non-union, however, a few mill hourly employees were unionized. At the completion of the layoffs. only a handful of employees remained for administrative and miscellaneous functions.For Getty employees, the terms of the merger agreement provided a generous level of separation benefits. The terms that were insisted upon by Mr. Petersen, Chairman of the Board for Getty Oil, stipulated that employees receive one month of severance pay, for each full calendar year of employment up to a maximum of two year's pay. This was the forerunner of what is now commonly known as the ""tin parachute”. All but the union employees of Petrotomics received these separation benefits. Union employees were covered by their separate, less generous, contract.Much of the literature that has been published in recent months has centred around the effects of layoffs on remaining employees, or the psychological effects on the severed employee. This study will emphasize the economic impact and ramifications due to a layoff, using the Petrotomics employees as a case example."

Jan 1, 1989

By Neil Schofield

"The estimation of recoverable ore reserves for feasibility in precious metal deposits such as Pine Creek must account not only f or the extreme variability of the ore grade but also for the complexity of the geology. Compatible models f or both must be adopted. Further, very selective mining demands reserve estimates be based on small selection units whose individual grades cannot be estimated reliably from the exploration drilling data.Multigaussian kriging is one method based on a multivariate normal model for estimating recoverable reserves under such conditions. In this paper, a non-mathematical description of the multigaussian model is presented together with its application to the Pine Creek Gold Deposit. The results, when compared to the first year of production, are encouraging for the model but not without surprises.Together, Part 1 and Part 2 of this study combine a detailed and geologically conditioned variogram analysis with an appropriate estimation method for a highly variable low-grade gold deposit. This and other studies of gold deposits familiar to the author show convincingly both that while the appropriate geostatistical skills and techniques are necessary to tackle the question of reserves, equally important but often discounted is the need for an appropriate and compatible geological model. IntroductionThe Pine Creek gold mineralization is associated with saddle reefs, and quartz and quartz- sulphide veining within the Enterprise Anti cline. The anticlinal structure is complex and, prior to mining, the mineralization was considered to be mainly strata controlled. Grades vary up to 160 grams in l .5rn core samples with coefficients of variation in excess of three.The estimation of the recoverable reserves for such a depo sit must account for both the geological complexity and the extreme variability of the gold grade.Part 1 of this work presented the geological model for the deposit and summary statistics 0 f the gold data. A detailed structural and variogram analysis considered both gold grades and the normal score (or gaussian transform) of these gold grades for different composite sizes."

Jan 1, 1988

By Robert J-M Ethier

"Sequential blasting has been introduced at Mines Gaspe to reduce the harmful effects associated with the instantaneously primed blasting method. Three stages of modifications had been tried to improve this method: trials of a rectangular pattern instead of the square one, trials of lengthened hole collars at the perimeter of the blast and trials of increased inter-row delay intervals from 35 to 50 ms. Results showed that improvements to the instantaneously primed blasting method were limited by the probability of increasing the number of bootlegs.Doing these modifications allowed the definition of a general approach to improvements before the sequential blasting method was introduced. This general approach included: planning of the mining sequence, following up on the drilling, loading and tying-in of each blast, compiling and evaluating blast results, comparing results from one bench to the other and moving along with one change at a time. This approach has permitted the successful introduction of the sequential blasting method in two steps: trials of blast by rows of delayed holes and trials of hole-by-hole blasts. The optimum result has been analyzed by high-speed camera and by computer blast simulations, The actual optimum sequential blast is obtained with a CORDLINE priming system using period 15, permitted in a large drilling pattern, with an intersequences delay interval of 230 ms and an inner-sequences delay interval of 50 ms.While these modifications were being introduced, a careful study of the drilling and blasting costs was made, which showed a sharp reduction of drilling cost per tonne. Even though there was an increased use of packaged slurry the blasting cost per tonne also dropped. The reduction in cost per tonne was explained by the fact that the average energy factor was maintained while the blasting pattern was expanded.The sequential blasting method, successfully introduced at Mines Gaspe, permits blasts of better quality, reduces the drilling requirements by more than 30% and helps to achieve improved productivity while minimizing costs."

Jan 1, 1984