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By J. Frostiak

"This paper describes some energy reduction measures Implemented at the Griffith Mine. Before proceeding, I will give you a brief overview of the operation.The Griffith Mine is an iron ore mining, concentrating and Pelletizing complex located in the red lake area of northwestern Ontario. The operation is owned by Stelco Inc. And is managed By Pickands Mather & Co. of Cleveland, Ohio. Operation of the concentrator commenced in 1968 and it has been producing at a nominal capacity of 1,500 ,000 tons annually.The plant flow sheet is illustrated in this slide. Run of the mine ore is crushed in a single stage 54"" gyratory crusher. Reclaim ore feeders draw from an open stockpile to feed two 32' diameter by 12' long autogenous mills, each driven by 6800 hp. Two s·rages of wet magnetic separation, roughers and cleaners, and primary cyclones complete the primary circuit. Cleaner concentrate is then ground in four 14' diameter by 28' long pebble mills rated at 2000 hp. The secondary grinding circuit is closed by cyclones and fine screens. Secondary cyclone overflow is further upgraded through hydro-separators, finisher magnetic separators and flotation.The magnetic concentrate produced is then thickened, filtered, balled and fired in three vertical shaft furnaces to produce a high quality hematite pellet."

Jan 1, 1983

By S. McAlister

"Falcon Concentrators Inc. has developed a high capacity centrifuge for the recovery of fine, free gold as well as other metals and minerals which were previously thought to be unrecoverable by gravity means. Development started in the early 1980's and the first commercial operation was commissioned in 1986. Gold applications include scavenging conventional gravity and cyanidation circuit tailings, cleaning gravity concentrates, cleaning flotation concentrates to melting grade, scavenging flotation tailings, and placer mining. Recent development work by Falcon has expanded the applications scope of the centrifugal concentrator to include many minerals amenable to gravity concentration including iron, tin, industrial minerals, and hazardous wastes. This paper discusses this technology including its history, present applications, and future potential. OVERVIEW OF CENTRIFIJGAL CONCENTRATIONUsing centrifugal force to enhance the separation of materials by differential specific gravity has been researched for over a century. Many attempts have been made to concentrate minerals in high gravity fields. Patents were granted on this subject as early as 1891 (see Figure 1). Patenting activity continued into the 1920's and has been relatively quiet since. It is significant that no centrifugal system of gravity concentration has been broadly accepted by the mineral processing community until recently. Much of the world's reserve of mineral deposits which are amenable to conventional gravity concentration technology (coarse-grained) have been mined and therefore there is an increasing demand for technology which can economically recover the fine-grained gravity recoverable minerals in the remaining deposits. In addition, centrifugal concentration can play an increasing role in solving environmental problems by offering an inexpensive method of concentrating hazardous materials such as uranium, plutonium, mercury, lead, and other heavy metals. The overall cost of disposing of these substances safely can be drastically reduced if they are recovered into a concentrated form prior to disposal."

Jan 1, 1992

By Thorndycraftm R. B.

"PURPOSEThe purpose of this presentation to the Canadian Mineral Processors is to review gold heap leaching systems currently in use and discuss their applicability to Canadian ores. The author's experience with the development and operation of heap leaching spans the past four years at. Pinson, and visits to many ether heap leaches in the western United States.INTRODUCTIONBy way of introduction to the heap leaching system, I will describe in slight detail the method used at Pinson Mining in northern Nevada. This will be followed by a discussion of optional methods in use at other operations visited.The Pinson orebody is a finely disseminated, well-oxidized hydrothermal deposit. Gold particles are finer than 20- microns. The orezone hostrocks are metamorphosed and silici¬fied limestone and shale. Sulfides and carbonaceous materials are negligible.The Pinson open-pit mine has operated since June 1980. It produces 1500 tpd of C.I.P. mill feed averaging .10 oz gold per ton, and 1500 tpd of heap leach material averaging .025 oz gold per ton. Cutoff grades are 0.04 for mill feed, and 0.01 for heap leaching. Stripping ratio is 9 tons of waste and leach grade to 1 ton of mill feed.The project is located in northern Nevada at 4900 feet above sea level.,Winters are fairly severe, with freezing conditions affecting operations from early December through April. Although officially designated as ""high-mountain desert"" country, snow and/or freezing rain occur almost every week of the winter.To date, all Pinson production-scale heap leaching has treated run-90,000 tons of ore are stacked on 300 foot square clay leach pads. Cyanide leach solution is applied through ""wobbler"" sprinklers at .0025 gpm/sq ft for 60 to 90 days. The pregnant solution flows from the heap by gravity to plastic-lined ditches, into a collection sump and then through buried piping to the 2.2 million gallon pregnant solution pond."

Jan 1, 1984

By James A. Finch, Stéphanie Somot

"The self-heating facility developed at the Noranda Technology Centre [Rosenblum and coworkers, 2001; 1995] has now been established in the Department of Mining, Metals and Materials Engineering at McGill University. A research program into mechanisms has been initiated (with support now from four industrial partners under Camiro). The first investigation is into the role of pyrrhotite (Po), the most commonly suspected mineral to trigger self-heating. Following a standard experimental protocol, self-heating was assessed on samples of Po mixed with sand as an inert material. With increasing Po content, self heating progressively increased but, interestingly, the samples visually appeared less and less oxidized. A high Po content, it was hypothesized, may give a more reducing environment favouring formation of H2S. The presence of this gas could be significant as it has a high exothermic heat of reaction with oxygen. The system was modified to test the hypothesis by including copper pieces and copper sulphate to react with (and therefore reveal) the gas as copper sulphide. Evidence is presented in support of H2S production, which to our knowledge has not been shown before and introduces a potentially important avenue of investigation into self-heating mechanisms.INTRODUCTIONSelf-heating of sulphide minerals poses potential environmental and safety hazards in mining of ores, and storage and transport of concentrates. While some conditions are known to be critical, the reactions remain poorly understood making self-heating difficult to predict and control.A major advance in measurement was introduced by Rosenblum and co-workers [Rosenblum et al., 2001; Rosenblum and Spira 1995]. A self-heating assessment method was developed and applied to various sulphide-bearing materials The method is based upon automated calorimetric cells with data acquisition, now available to McGill researchers and which have been used for the study reported here."

Jan 1, 2006

By Louis J. Cabri

"Mineralogical evaluation of ores containing platinum-group elements (PGE) is fraught with difficulties because of the low grades (often 1-3 g/t PGE), dispersal of PGE (usually as finegrained platinum-group minerals [PGM]), as well as occurring in solid solution in some base metal sulphides (BMS) such as pentlandite and pyrrhotite. Mineral processors not only need to know how the PGE are distributed, that is, their mineralogical distribution, but the grain sizes of the PGM, and especially the associated minerals.After obtaining a representative ground sample, common approaches include studying a large number of polished or polished thin sections, concentration of the PGM and sulphides by different gravity methods (e.g., heavy liquids, elutriation, micro-panning) and study of the produced concentrates. Unfortunately, these methods often do not produce concentrations that are sufficiently representative of a particular ore. The newly developed patented technology of hydroseparation (Rudashevsky et al., 2002) is especially effective at making concentrations of PGM over a wide range of concentrations (down to <300 ppb Pt+Pd+Au) and size fractions, including <45µm (down to a few micrometres in size), a size range not possible by other techniques. This water-based environmentally friendly laboratory technique provides a large amount of data from samples, which can range from 5 to 1000g, giving confidence to extrapolation of results to bench-scale and larger test work, especially when combined with quantitative image analysis to determine modal percent of major minerals and mineral liberation data, needed to predict the minimum (K80) and optimum (K40) grinds. INTRODUCTIONThe invention of the electron microprobe in the early 1960’s and scanning electron microscopes (SEM) with energy dispersive spectrometers (EDS) in the 1970’s presented mineralogists with powerful tools permitting them to obtain chemical and surface area data on minute amounts of minerals in situ, in a non-destructive manner. An area of process mineralogy that has very much benefited from this new technology was the characterization and understanding of PGM species, resulting in rapid growth of newly characterized species (Figure 1). Recognition of the first PGM may be ascribed to the Spanish explorer Ulloa in 1748, who described native platinum from South America. The second PGM (laurite, RuS2) was only recognised in 1866 by Wöhler in placer samples from Borneo (now Kalimantan) and the third (sperrylite, PtAs2) by Wells in 1889 from the Sudbury Ni-Cu sulphide deposits. The discovery of the important Pt-Pd deposits in the Bushveld Igneous Complex led to the characterization of the important Pt-Pd sulphides (cooperite, PtS2 in 1928 and braggite, (Pt,Pd)S2 in 1932). By the early 1960’s, only about 20 PGM were recognised, and many later proved to be poorly or incompletely characterized. Thus, the introduction of electron probe microanalysis (EPMA) led to a 50% increase of known PGM, as shown in a 1972 review on PGM (Cabri, 1972). At the time of the next review in 1981 (Cabri, 1981) the number of known PGM had grown to 75 and as of 2002, it had increased to 109 (Cabri, 2002). However, it is the author’s experience that less than about 25 PGM may be considered to be “common”, i.e., found in many different deposit types, as shown in Table 1, but usually only some of these occur in any one particular deposit."

Jan 1, 2004

By D Reddy Chinnasane

Outline Location of BirchtreeMine and Layout of 124 OB Cut and Fill Mining Background Purpose of the Studies Methodology Results and Analysis Conclusions and Recommendations

Oct 1, 2010

By P. Toscano

The objective of this investigation was to determine how the losses of nickel, copper and cobalt in fayalite slags could be controlled for nickel mattes with iron contents below 5 wt%. Laboratory-converting and slag-cleaning experiments were carried out at 1250°C. Several variables were investigated during and after converting to help minimize the amount of nickel and cobalt that reports to the slag. The variables used were oxygen concentration in the blast, coke breeze addition and nitrogen injection stirring. In all the experiments, slag and matte co-existed. Coke additions during converting have little effect on the losses of nickel, copper and cobalt. In fact, coke additions retard the converting process by consuming some of the oxygen injected. On the other hand, coke additions and nitrogen injection applied after converting was completed resulted in decreased slag losses.

Jan 1, 2004

By P. Toscano, T. A. Utigard

"The objective of this investigation was to determine how the losses of nickel, copper and cobalt in fayalite slags could be controlled for nickel mattes with iron contents below 5 wt%. Laboratory-converting and slag-cleaning experiments were carried out at 1250°C.Several variables were investigated during and after converting to help minimize the amount of nickel and cobalt that reports to the slag. The variables used were oxygen concentration in the blast, coke breeze addition and nitrogen injection stirring. In all the experiments, slag and matte co-existed.Coke additions during converting have little effect on the losses of nickel, copper and cobalt. In fact, coke additions retard the converting process by consuming some of the oxygen injected. On the other hand, coke additions and nitrogen injection applied after converting was completed resulted in decreased slag losses."

Jan 1, 2004

By K. D. Mills, R. G. Robins, C. G. Anderson, L. G. Twidwell

Arsenic is a metalloid which has attracted great environmental scrutiny in recent decades. Moreover, its control and deportment is becoming increasingly important in the metallurgical processing of modern complex ores and concentrates. As such the authors of this publication have a combined professional experience of well over 100 years in the study and hydrometallurgical treatment of arsenic. Hence, this paper will cover arsenic fundamentals along with a review of technology for removal, fixation and stabilization. Finally, an application of industrial Alkaline Sulfide Leaching (ASL) of a copper enargite concentrate will be elucidated.

Jan 1, 2014

By Y. Shu, A. R. Laplante

"A laboratory Knelson Concentrator (LKC), 7.5 cm in diameter, was used to study the performance of a similar unit at plant scale, 76 cm in diameter, This paper reports on sampling and assaying procedures and problems, and discusses the performance of the plant unit.The sampling procedure was based on the extraction of large samples, 5-15 kg, and their complete processing with a LKC, to characterize size-by-size performance. The procedure yielded reliable estimates of overall gold assays, after blatant assaying errors were corrected. The plant centrifugal unit concentrate was sampled in situ, which yielded very erratic grades, with a large overall negative bias. Due to the very heterogeneous nature of the concentrate, which can be linked to flow patterns in the riffles of the unit, it is recommended that for future testwork, the concentrate be recovered separately, blended and then sampled.The plant centrifugal separator, when operated at a feed rate of 25 t/h, yielded total and free gold recoveries of 41 % and 62%, respectively. This is much lower than previously reported data, and suggest that at 25 t/h, the unit is overloaded. However, recovery did not decrease with time over a loading cycle of 2 hours, which suggests that present loading cycles, generally 1 hour, can be significantly increased. INTRODUCTIONThe study of gold gravity circuits is plagued with the statistical problem of determining the behaviour of particles occurring with a very low frequency. This problem becomes particularly acute with increasing particle size. For example, considering the average weight of a gold flake in the 840-1200 µm (20/28 mesh) to be about 5 mg (Banisi et al, 1991), a reasonably high grade feed of 10 git would require, for a relative error (standard deviation) of 10% on grade, 100 particles, that is 0.5 g of gold, or 50 kg of 840-1200 µm material (a Poisson distribution is assumed). Gy's formula, for the same data, yields a weight of 33 kg, with a shape factor of 0.1 and a size distribution factor of 0.5. Both results are comparable; to obtain the suggested weight of material, a stream containing 5% weight in the 840-1200 µm would require 1 tonne samples (of unsized material). The problem rapidly disappears with decreasing particle size. Still with Banisi et al's data, a gold flake in the 300-420 µm range averages 0.5 mg in weight; the same 10% error on grade corresponds to a weight of 5 kg. Gy's formula for the same size distribution and particle shape factors as above predicts a required mass of 1.4 kg. This would correspond to an unsized sample weight of 10 kg with 14% weight in the 300-420 µm, for example."

Jan 1, 1992

By R. Grant, C. Bazin

"Brunswick Mining has investigated grinding media optimization for the purpose of improving the flotation feed and regrind product size distributions to enhance metal recoveries. This paper presents the findings of on-going testing of various sizes and shapes of grinding media. Plant observations combined with estimated breakage rates demonstrate the importance of appropriate selection of media sizes on the milling operation.IntroductionBrunswick Mining and Smelting is located near Bathurst, New Brunswick, and operates one of the world&apos;s largest lead and zinc mines. The ore mineralogy and processing flowsheet have been described elsewhere (Petruk and Schnarr, 1981; Shannon et 01., 1993). Despite the 25 years&apos; milling experience, there has been considerable disagreement as to the required grind size (Grant et 01., 1991). Since 1964, fineness of the flotation feed ranged from 60% to 80% passing 400 mesh (37 um) and averaged slightly above 71% in 1992. Results from mineralogical analysis (Petruk and Schnarr, 1981), laboratory and pilot plant experiments, as well as plant sampling (Grant et al., 1991; Hendriks and Ounpuu, 1985; Shannon et al., 1993) indicate that improved metal recoveries can be achieved by an increase in the fineness of grind, and more specifically by maximizing the amount of material in the 15 urn to 25 um range.This paper shows the results obtained so far from the evaluation of the effect of grinding media size and shape on grinding mill efficiency. The problem of finding optimum grinding media size, shape and composition has been recogrtized since the early 19405. The main criterion for the selection of media size is the size of the material to be ground (Bond, 1958). Although several relationships have been proposed to assist in the selection of grinding media size, none of them are entirely satisfactory (Wills, 1988). Other approaches using the breakage rates and process simulators were developed to help in the selection of grinding media size (Austin et al., 1976; Lo and Herbst, 1986; Concha et 01., 1992; Samskog et al., 1990; Martinovic et al., 1990). Although considerable effort was put into the analysis of grinding media size, very little information is available on the effect of grinding media shape (Norman and Decker, 1985, p. 3C-4). Norris (1952), Kelsall et al. (1973) and Herbst and Lo (1989) have shown that balls are more efficient than other shapes of grinding media including truncated cones and cubes."

Jan 1, 1994

By Norman O. Lotter, Dominic Fragomeni, Lori J. Kormos, Guy Comeau, Antonio Di Feo

"A performance opportunity was identified for the Raglan concentrator operation by changing the selection of xanthate from Potassium Amyl Xanthate (PAX) to Potassium Isobutyl Xanthate (PIBX). This opportunity was identified by a Statistical Benchmark Survey of the operation and QEMSCAN measurement of these survey samples, in which fine liberated pentlandite was found in the final tailings. This was confirmed by some laboratory scale flotation testwork using High- Confidence Flotation Testing. A statistically designed plant trial was implemented in 2007 to measure and confirm the performance gain. This trial design used on-off switching following a three-month reference distribution of normal operations using PAX. This layout provided both external and internal reference distributions on PAX to be established. These were used to evaluate the internal reference distributions on PIBX. This trial format was deliberately designed so as to block out the effects of autocorrelation in the operations data. In this way, small but significant metal recovery gains may be measured and proven in a plant trial. The paper describes the case study to show that this approach successfully blocks out autocorrelation in time, which is an obstacle in analysing concentrator operations data. The results showed that the PIBX had produced clear and statistically significant recovery gains for Ni, Cu, Pt and Pd. Adjusted for feed and concentrate grades, these recovery gains were: Ni: 1.00, Cu: 0.96, Pt: 5.07 and Pd: 1.90%.INTRODUCTIONConcentrator OperationsMilling and flotation, sometimes accompanied by gravity separation, are well-established unit processes. These have traditionally been used to treat ore that has been mined, producing a lowbulk concentrate containing a high percentage of the paymetal(s) in the ore mined. These concentrates are more economically treated in subsequent beneficiation processes than the ore from which they were produced.This choice of a concentration process results in a significant and variable paymetal loss to the concentrator tailing. This tailing is discarded and disposed of in an impoundment area, and has no immediate business use. This paymetal loss forms the largest loss of paymetal in the entire flowsheet by which refined, or saleable, metal(s) are produced [Cramer, 2001]. The question arises as to whether any particular concentrator is performing at its full potential, and how this answer may be determined. In a concentrator operation, there is always an ongoing performance improvement programme for this reason."

Jan 1, 2009

By N. Mejia, Z. Xiao, D. Fragomini, A. R. Laplante

"The adaptation to platinum-group minerals (PGMs) of a protocol developed to predict gold recovery by gravity is presented. The protocol is based on three different data sets: ore characterization of gravity-recoverable minerals, description of their behaviour in cyclones and ball mills, and performance of recovery units. Examples of these are given. It is shown that the gravity-recoverable PGM (GRPGM) content can be very high. A case study at the Clarabelle mill is presented. Sperrylite is the dominant platinum mineral, and its partition curve is very similar to that of gold, despite its lower density. Its grinding kinetics is significantly higher than for gold. As a result, sperrylite accumulates in finer sizes than native gold, but can still be recovered by gravity or flash flotation. Recovery projections based on gravity recoverable data are presented.BACKGROUNDReliable modeling of gold recovery by gravity or flash flotation from a grinding circuit is based on the ability to describe the properties of gold-bearing particles in the ore, their behaviour in grinding and classification units and the performance of the equipment used for recovery, be it gravity or flotation. This is shown in Figure 1."

Jan 1, 2004

By Fernando Romero Lage

The purpose of this paper is to analyze detailed crushing and screening configurations and the interrelationship as they relate to mass flow balance, operating efficiency, plant availability and final production. The entropy of ore bodies limits the standardization of crushing and screening plants, however, consideration for a continuous and sensible assessment of the mass flow balance of the circuit will yield the best achievable results. Since energy costs have continuously risen in the last few years, it is imperative that these circuits offer low power requirements and low operating expenses. The paper will detail how the efficient implementation of continuous mass flow balance in the application of crushing and screening plants will generate quality improvement and process improvement, which will translate into bigger production and energy savings.

May 1, 2010

By A. H. A. Robinson

Until quite recently-practically up to the present time-all but an insignificant amount of the copper produced in Canada was shipped abroad in the form of ore, matte, blister, etc., there to undergo the further treatment necessary to bring it to the refined state in which it is required by fabricators of copper. Consequently, Canadian manufacturers were forced to import nearly all the copper required in their business; and a considerable portion of the ultimate value to be derived from a Canadian raw material was lost to the country of its origin. In the early days of Canadian copper mining, when both the production and consumption of copper in Canada were much smaller than now, such a condition of affairs was perhaps economically unavoidable, and, in any case, was not a matte~ of sufficient importance to attract general attention. Later on, however, when there had been large and rapid increases both in the production and consumption of copper in Canada, it formed the basis of much adverse criticism, deserved or not, of producers of Canadian copper-criticism that became intensified during the war and subsequent years. Happily, the chief grounds of complaint have now been removed, and there are today copper refineries in Canada, either operating or under construction, that are capable, in the aggregate, of turning out amounts of refined copper greatly in excess of any indicated Canadian consumption.

Jan 1, 1931

By V. I. Lakshmanan

Process Research ORTECH Inc. (PRO), with its in-depth knowledge on cobalt behaviour in nickel matte smelters and its novel mixed chloride hydrometallurgical processing, is developing a hybrid process to enhance cobalt recovery. PRO’s experimental data is compared with plant information on cobalt behaviour in smelters. In final stages of iron rejection in the converters, the thermodynamics of the matte-slag system dictates the undesirable transfer of the cobalt to the final converter slags. This results in much lower cobalt recovery in the product Copper/Nickel matte (<60%) in present copper and nickel smelters. This information is used to transfer cobalt into desired quantity of converter slag. The enriched cobalt slag is subjected to PRO mixed chloride process to recover nickel, and cobalt as high value salts of importance in lithium ion batteries. From a typical converter slag containing 3.5% Ni, 1.5% Co, and 1.0% Cu, PRO leach process extracts 98 % Ni, and 98 % Co, and 99% Cu. Application of this process to enhance cobalt recovery in typical Copper/Nickel smelters is presented. In addition, this will also enhance the Copper and Nickel recovery. Both these aspects are of interest to the Copper, Nickel, and cobalt producers. INTRODUCTION In pyrometallurgical extraction of copper and nickel from sulphide ores cobalt often occurs with copper and nickel. The requirement of cobalt is increasing due to its need in lithium ion cells used in electric vehicles. Therefore its efficient and economical extraction in a form suitable for use in lithium batteries is of great interest at present. In the pyrometallurgical processing of copper and nickel sulphide concentrates the concentrate is smelted to produce a furnace matte, which is then sent for converting to remove the iron to produce iron free (~1% Fe) matte for further processing to copper or nickel matte. The copper/nickel matte is processed for metals recovery by hydrometallurgical or vapo-metallurgical or oxidation – reduction routes to make market products. As shown in Figures 1 and 2 the slags are produced in the smelting step and in the converting steps. The metal losses in smelting are controlled by the oxygen pressure in that step and slag production which is controlled by concentrate grade. The slag produced in the converting step which is carried out at a higher oxygen pressure contains higher level of metals is often returned to the smelting step to increase value metals recovery.

Jan 1, 2019

By Gabka T., R. E. Gregorczyk

Work to improve employees' behavior in Głogów Copper Smelter began in 2009. It has been noted that over 90% of accidents at work are attributable to human reasons. At that time, a series of external workshops was contracted based on exercises conducted by a psychologist trainer with supervisors. Having trained about 300 employees, an observation of employee behavior at workstations was started as part of the control of risk reduction measures. As a result of the above actions, the number of accidents was reduced by half between 2009 and 2017. Currently, internal workshops for communication improvement for behavioral auditors are conducted by the employees of the Smelter. Named: "I observe, react and communicate" and aimed at further mastering the elements of the safety culture in Głogów Copper Smelter. INTRODUCTION KGHM Polska Miedz is an integrated mining and smelting copper producer based in Poland and operates on the international metal market. KGHM, with annual production of over 500 thousand tonnes of copper, holds a middle position in the world ranking of copper producers. At the same time, in Europe it is the largest producer of copper from its own ore deposits, possessing its own integrated production structure, from underground mining operations to final, top-quality product. It is also worth mentioning that KGHM Polska Miedź is one of the largest Polish exporters and, apart from copper, is a significant primary silver producer in the world. The mission of our company, consistent with its strategy, is to produce copper and silver in a safe mode for workers and the environment, with reduced environmental impact. Therefore, "Głogów" Copper Smelter operations are continuously focused on the creation of safe working conditions, and minimizing the negative effects on the environment. The company meets the requirements of an Integrated Management System comprised of the following: Quality Management System ISO 9001:2008 Environment Management System ISO 14001:2004 OHS Management System according to the requirements of BS OHSAS 18001:2007 Being a proud socially responsible company, we are aware that part of an efficient OHS system is to recognize, monitor, prevent, educate and motivate, thus continuously improve working conditions for 2500 employees of Głogów Copper Smelter &Refinery. We clearly communicate that a low accident rate is beneficial for both, staff and the company, and is of utmost importance as a part of our health policy.

Jan 1, 2019

By Carol E. Jones

Outline ? Reclamation is a primary mitigation. ? Understand baseline environment and land use. ? Determine reclamation goals. ? Selecting reclamation technology. ? Determine technical risk. Reclamation ? Primary Mitigation ? Through land and water disturbance, mining can have significant local and regional environmental impacts. ? Impacts can also result from associated land use changes. ? Reclamation is a primary tool that the mining industry uses to manage its environmental impacts.

May 1, 2008

By Kjell Winther

"Hydralift Scanmec introduced early in the eighties a new and different mill design for very large grinding mills. The design described has actually no size limitation upwards and is so far suggested with motor power up to 40 MW. The mill is furnished with gearless ring motor drive transmission. It is totally fabricated, without heavy castings. The first installation, an 8, 1 MW ball mill dia. 6,5 m x 9,65 m length, has been in continous operation in 13 years without problems. Total grinding costs has been reduced with 3 0% compared to traditional smaller mills. Reductions has been achieved in maintenance cost, power consumption, liner wear and grinding media consumption.So far 3 ball mills of the described design are delivered and in operation, one grinding Iron ore and two grinding Nickel ore.IntroductionThe design of mills for mineral ore dressing has basically been unchanged since grinding in mills started some I 00 years ago, in spite of the fact that mills have grown larger and larger. This has resulted in structural and mechanical problems, causing expensive failures and loss of production.Hydralift Scanmec engineers were designing mills for more than 20 years according to the old «rules». As they experienced interest for an increase in the mill size at the time when reports about failing mills were published, they started analyzing the different mill parts to find new solutions. One goal was to only use elements from proven technology in the new design. Two key items were ring motor (wrap around motor)(Isch et al. 1977, Meintrup and Klein 1982, Ranze 1983) and large diameter pad bearings (wrap around bearings)(Fuller 1984), both acknowledged from other industries."

Jan 1, 1995

By Leslie F. Sawatsky

Failure of drainage works is one of the principal threats to the sustainability of mine closure facilities but the risk can largely be remedied by appropriate design of mine closure drainage facilities. The feasibility of sustainable mine closure facilities that avoid failure of drainage systems needs to be established during the early stages of mine planning, before mine disturbances make sustainable closure uneconomical. Experience in mine development shows there is often enough latitude in mine plan development to incorporate measures with neutral economic consequence and positive environmental benefit. Project managers need to resist the temptation to develop mine plans and closure drainage plans independently. Integrated mine planning with effective inputs by environmental specialists and drainage designers may affect the overall mine configuration, footprint and schedule. Key issues pertain to the end pit location, waste dump configurations, pit infill elevations, fluid impoundments, diversion channel routing and preservation of flows to receiving waters. Guidelines with illustrations of suitable practices are presented to assist mine planners.

May 1, 2004