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By Thomas M. Barczak

A critical safety component for all underground mining is intrinsic and standing support systems. Specifically, tabular hardrock mining exploits a wide variety of deposits that require various methodologies to safely mine the ore material. Several new support systems have been developed in recent years for hard rock applications which have been modified or applied in a variety of ground control applications. These include prestressing devices, improved cribs, and free standing supports. Various standing support and prop-type systems have been designed for hard rock applications with seismic loading conditions to accept prestressing. The prestressing, using water-filled cells, creates an active load upon installation and is considered essential to maintain support during and after the blasting of the mine faces. Heavy seismic activity is also present in these mines and the prestressing units can provide some energy absorption capability to help preserve the integrity of the support. This paper presents an overview of these innovations in support technology.

May 1, 2007

By Carolyn Kenney, Jianjun Liu, Zhenghe Xu

"The flotation of fine particles has been a challenge to scientists and engineers in mineral processing. A novel concept of reactive oily-bubbles as carrier to float fine particles is described in this paper. The reactive oily bubble flotation is considered attractive for fine particle flotation. In addition to the role of fine particle agglomeration by oily films, the surface properties of air bubbles coated with a thin oil film can be better controlled for the desired selectivity by adding certain types and concentrations of water insoluble collectors into the oil phase. This approach can eliminate the addition of collector to the aqueous phase, avoid undesired synergetic interactions among collectors, activators, depressants and dispersants present in slurry, minimize undesired activation of gangue particles and significantly reduce the amount of collectors needed.The concept of using reactive oily-bubble to achieve selective flotation was demonstrated in contact angle measurement and by microflotation tests with fine sulphide minerals: sphalerite and galena in this paper. The flotation behavior of sulfide minerals can be radically modified with adjustment of solution pH, addition of depressant and activator as in conventional flotation. The successful separation between PbS/SiO2, ZnS/SiO2 and PbS/ZnS was demonstrated using mixed mineral systems. Compared with conventional flotation, the reactive oily-bubble flotation showed a stronger collecting power.INTRODUCTIONThe challenge for fine particle flotation has long been recognized and well documented (Fuerstenau, 1980). Because of drastic changes in surface properties of fine particles and associated physicochemical and hydrodynamic conditions of flotation systems, fine particles responded poorly to the conventional flotation processes. For instance, the small mass and hence low inertia of fine particles would result in a low efficiency of collision and attachment of fine particles with flotation bubbles. Large specific surface area of fine particles, on the other hand, would cause an increased reagent consumption. Dominating colloidal forces of fines may cause slime-coating, accompanying a low selectivity (Somasundaran & Arbiter, 1979; Fuerstenau, 1980; Sivamohan, 1990)."

Jan 1, 2003

The importance of understanding the behaviour of cemented paste backfill (CPB) within the framework of the effective stress concept is highlighted in this paper. Only through an understanding built on this concept can problems such as barricade loads, the rate of consolidation and arching, be fully understood. The critical importance of the development of stiffness during the hydration process, and its impact on factors such as barricade loads, is used to illustrate why conventional approaches that implicitly ignore the effective stress principle are incapable of capturing the essential components of CPB behaviour.

May 1, 2007

By D. Fleury

"This paper describes the challenges of mining a sill pillar under various types of paste backfill quality. From 1994 to 2002, Louvicourt mine produced approximately 12,600,000 t grading 3.5% Cu, 1.5% Zn, 27 g/mt Ag, and 0.85 g/mt Au. The mining method is transverse blasthole stoping, mining primary and secondary stopes. The production of this 4,300 t/d operation was accomplished using two mining horizons; from the 655 m level to the 415 m level and from the 860 m level to the 680 m level. The main sill pillar is between the 680 m level and the 655 m level. The second sill pillar is between the 885 m level and the 860 m level representing a quarter of the main sill pillar in size. Since 2002, production has gradually decreased and mining activities are expected to end in mid-2005. The sill pillar production will represent almost 25% of the total mine production. The overall recovery of the sill pillars stopes will be discussed in this paper. To recover the sill pillars, more than 1,500 m of development was planned in or under paste backfill. The first sill pillar stopes under paste backfill were mined out in 2002 with excellent results."

Jan 1, 2005

By Son Hye Jin

Abstract ? KHNP is participating in the construction of Nuclear Power Plant both in Korea and UAE. ? We want to improve our construction project management process. ? So we are now developing the process model on nuclear power plant construction management. ? For the effective adoption of this process, we are planning to develop IT system. ? For the first step, we were established the Information Strategy Plan

Aug 1, 2013

By Robert E. McIvor

"“Classification System Efficiency (CSE)” is a powerful, emerging tool that can be used to increase grinding circuit efficiency. It is defined as the percentage of the ball mill energy being used to grind target, oversize particles. The discovery, validation and current use of CSE, including how it was manipulated, are described through seven plant case studies. These resulted in efficiency improvements ranging from 3 to 30 percent, along with major reductions in over-grinding. The achievable CSE for any given circuit is described. This paper demonstrates both how, and how much, every plant can benefit from increasing grinding circuit CSE.DEFINITIONS: BALL MILL CIRCUIT “CLASSIFICATION SYSTEM EFFICIENCY (CSE)”, “EFFECTIVE MILL GRINDING POWER”: “WASTED/OVER-GRINDING MILL POWER”The closed ball mill circuit is shown in Figure 1. The extreme importance of the circulating load ratio in determining circuit productivity was demonstrated by Davis (1925) and re-published by Guadin (1939), as shown in Figure 2. The circuit production rate increases with increased circulating load ratio, rapidly so at first and continuing to increase more gradually with higher and higher values of circulating load ratio.The underlying cause of this relationship was seen when comparing the ball mill feed and product size distributions from two different ball mill circuits, one with a low circulating load ratio, and one with a high circulating load ratio (Figure 3). While the F80’s and P80’s of these two circuits were close to the same, the mill feed and discharge size distributions were quite different. The ball mill in the circuit with the higher circulating load ratio contained a substantially greater fraction of material coarser than the target P80, as indicated by the respective mill feed and discharge size distributions. This fraction represents the fraction of the mill power being expended on “coarse” (in this case, plus P80 size) material, or the circuit “classification system efficiency.” The “effective mill grinding power” is the total mill power multiplied by the CSE. The compliment of CSE, the remaining fraction of “fines” in the mill, is the fraction of the mill power being wasted, and spent on over-grinding of material already finer than the circuit target product size. This is the “wasted/over-grinding mill power”"

Jan 1, 2013

By R. H. Pemberton

"THE Mattagami district of Quebec has, during the past few years, experienced intense exploration by virtually every mining company operating in Canada. As a result of this activity, a number of well-mineralized zinc-copper bearing bodies have been found, some of which will almost certainly go into production. Much of the area has undergone what is now termed ""Saturation Prospecting"" or intense and detailed geological and geophysical surveying, followed by the drilling of nearly all possible bets. Considering only the airborne geophysical surveying, some blocks of ground have been flown by virtually every geophysical instrument which has to date been installed in an aircraft. A list of these different geophysical instruments would include: l. Fluxgate-type total field magnetometers 2. Vertical field magnetometer ,3. Nuclear-precession magnetometers 4. Single-frequency out-of-phase electromagnetometers 5. Dual-frequency out-of-phase electromagnetometers 6. Horizontal component measuring electromagnetometers 7. In-phase and out-of-phase electromagnetometers8. Dip-angle measuring electromagnetometers."

Jan 1, 1961

By P. J. Gron

"The Samatosum Project is a base and precious metal deposit located in south central British Columbia. A dilute ore reserve of 766, 682 tonnes grading an average 1.1% copper, 3.0% zinc, 1.4% lead, 833 gram per tonne silver and 1.6 gram per tonne gold is available. The base metals occur as tetrahedrite, chalcopyrite, sphalerite and galena. Virtually all of the silver and gold are associated with the tetrahedrite.The mill is operated at a rate of 21 tonnes per hour. Sulphur dioxide is used as a zinc depressant in the grinding circuit. Production of a Cu/Pb/Ag concentrate is accomplished by conventional flotation. This concentrate is then separated into Cu-Ag and Pb concentrates through a reverse flotation circuit utilizing sulphur dioxide, starch and sulphuric acid as lead depressants.The tailings from the Cu/Pb/Ag circuit are treated with copper sulphate to activate the zinc. Production of a zinc concentrate is accomplished by conventional flotation. Dewatering of the concentrates to 6% moisture is done with a LAROX filter press.Tailings effluent is treated with lime and sodium sulphide to precipitate out the heavy metals. The resulting sludge is treated in OYNASAND sand filtration units to produce a quality effluent."

Jan 1, 1990

By Jan E. Nesset, Phillip J. Mackey, Joo Y. Kim, Giovanni Di Prisco, André R. Laplante

Understanding the appropriate gold recovery to associate with copper recovery when assessing laboratory testing on porphyry deposits is often complicated by several factors including variable and/or complex mineralogy, and the variability contributed by the low Au content of the various flotation products. As a result, gold recovery is often underestimated and the value contributed by the gold to project economics is less than it should be, a problem that could influence overall project viability. Typically, for these types of deposits, the gold recovery is found to be lower than the copper recovery. In this unique project, having the objective of seeking a better understanding of gold recovery relative to copper recovery, mill feed samples from a selection of six operating world-class Cu-Au low-grade deposits were collected along with the mill metallurgical results for the corresponding sample periods. The samples were subjected to laboratory flotation tests, as well as the Laplante gravity recoverable gold (GRG) test, mineralogical assessment and cyanidation leaching. The aim was to better understand the copper and gold recovery relationship, the links between laboratory and plant flotation performance, the influence of mineralogical differences, and the potential for enhancing gold recovery by gravity recovery techniques. A new correlation of copper and gold recoveries based on the results is presented in the paper. The results showed that the link between copper and gold recoveries established in the best laboratory tests were maintained in the plant results, and further, that the potential for improving Au recovery by gravity or flash flotation methods is directly proportional to the amount of gold greater than 75 microns.

Jan 1, 2005

By Donald Leroux, Colin Hardie

"New or improved mathematical models of mineral processing machines are published almost on a monthly basis. Often, these models can be calibrated from bench-scale test results and programmed using spreadsheet software to simulate different designs and operating conditions in open-circuit configuration. However, since mineral processing plants frequently operate comminution, separation and/or extraction machines in closed-circuit configurations, open-circuit simulations often have limited applicability.The simulation of closed-circuit configurations necessitates the programming of an iterative loop to bring the circuit to equilibrium. This can be done by using spreadsheet software but often requires advanced programming skills. As an alternative, LIMN® software, which simplifies the programming of closed-circuit configurations within spreadsheets can be used. This software was specifically designed to turn a Microsoft EXCEL® spreadsheet into a closed-circuit modeling and simulation platform with minimal programming.During the past few years, the Mineral Processing Group at the Noranda Technology Centre, has used this software for numerous projects which include the simulation of multi-stage cyclone circuits, closed-circuit regrinding and locked-cycle flotation tests. Using these projects as examples, the flexibility and the easy to use nature of the LIMN® software is demonstrated.INTRODUCTIONMineral processing engineers routinely have to make predictions regarding comminution and separation processes. Predicting the appropriate dimensions of equipment and the process response to changes in operating mode are probably the most common situations that they face. In all cases these predictions are expected to be rapid and accurate.Predictions made by engineers are frequently based on mathematical models. Many first principles, empirical, semi-empirical, steady state and dynamic models of mineral processes can be found in the literature. In fact, new or improved mineral processing models are continuously being proposed and in spite of this variety, process engineers often have to either develop their own or tailor existing models to suit their particular needs."

Jan 1, 2003

By Bryan M. Maybee

The scheduling of underground mining activities is a complex task, in particular because of the requirement to satisfy a large number of constraints. A Schedule Optimization Tool (SOT) for underground mine planning is described, and its use for investigating strategic decisions is illustrated by means of a case study. The main components of SOT include: a genetic algorithm, which searches for optimal schedules; guidance, the application of heuristic rules to guide the genetic algorithm; and sliding, which implements just-in-time development. For the case study, SOT is employed to generate schedules at various ore capacities, accounting for the appropriate financials in each schedule evaluation. The case study finds that preliminary scheduling performed before invoking SOT, intended to provide a good starting point for the search, does not ultimately lead to higher value schedules. Two key uses for SOT are identified. A mine planner can save time by using SOT to produce high-value scheduling alternatives. This process ensures a broader investigation of the value that the deposit can yield. Also, SOT can be used to validate strategic and tactical decisions, by scheduling under alternative scenarios. Both uses produce high value schedules conforming to all imposed resource constraints.

May 1, 2009

By W. R. Hargrave

"Mining OperationsOre will be mined by open pit techniques. Because of the high grade of the deposit and the low strip ratio, mining rates are small relative to most other open pit operations.The ore deposit is located in the valley of the main fork of Red Dog Creek. The bulk of the deposit is located on the southwestern flank of the creek, but mineralization does traverse the creek in both the north and east directions.Erosion has exposed the orebody in a number of locations, and has resulted in oxidation of the mineralization closer to the surface. Although much of this oxide material is very high grade, it leaches readily and, at least in the initial phases of the operation, will be stockpiled and not treated in the concentrator.During construction, 2.3 million mt. (2.5 million st.) of overburden and waste will be stripped from the orebody to expose ore for initial operations. Competent non-mineralized waste will be used to build roads and waste dumps, but the bulk of material removed consists of oxide or ice-rich soils and weathered rock which is not suitable for construction.The annual ore mining rate will be 1.9 million mt. (21. million st.). The strip ratio will start off at 1.2 to 1, falling to 1.0 to 1 by year 5 and ultimately to 0.8 to 1.Bench heights are set at 7.6 m (25 ft). Pit slopes will vary depending on the material present in the wall. The slope, through competent waste and the sulphide package, was set at 45 degrees. This is flattened to 26 degrees (i.e. 2: 1) through overburden and Upper Kivalina shales. The Kivalina Shales are an ice-rich shale whose carbonate matrix has been leached out and which exhibits little or no strength as it thaws. In addition, a 100 ft containment berm will be left in the wall below overburden and Kivalina."

Jan 1, 1989

By J. M. Menacho, J. M. Reyes

"Batch and continuous regrind tests were conducted in 3 kW and 15 kW tower mill units, respectively. Single particle size classes of a chilean copper sulphide concentrate were employed in the batch tests, while a fraction of the plant stream was deviated to the continuous unit operated under different conditions.Results are interpreted in terms of the simplified grinding kinetic model energy-based and conveniently adapted to this application. Based on the above figures plus some industrial data supplied by MPSI Inc., a scale-up relationship was developed for the apparent specific rate of breakage siE, as a function of the net power draw P, i.e.siE = siE text (P/P test) 0.089This equation fit very well the available data at the three scales and its application to regrind 4400 t/day of rougher copper concentrate from 70% below 325 Tyler mesh to 95% below same mesh, involves around 30% energy-saving compared to standard ball mill for same grinding task.INTRODUCTIONImpact and abrasion-attrition breakage are two extreme mechanisms present in any grinding system in a proportion defined by specific design and operating conditions. Impact breakage is generally efficient for coarse grinding, while abrasion-attrition is adequate to produce fine particles. Efficiency of conventional tumbling mills decreases as fineness of the product increases, even though distribution of applied stresses is always much stronger than internal force distribution in the particles. On the other hand.in a vertically stirred ball mill like the tower mill, distribution of applied stresses and particle strengths are in the same order of magnitude, making possible nearly ideal abrasion conditions (1).Tower mill configuration includes a cylindrical body vertically mounted and provided with a central screw with flights to produce circulation of balls moved up through the inner region to return down by the external space. Fresh material can be fed into the top , center or bottom in order to assure appropriate agitation of suspended particles. Size reduction and evacuation rates should compensate settling of coarse particles to avoid sanding of the screw. Detailed description of the tower mill and its operation have been given elsewhere (2),(3)."

Jan 1, 1989

By Carl Bertrand, Patrice Nadeau, Claude Bazin

"Rio Tinto Fer et Titane processes the ore extracted from the Lac Tio orebody to concentrate the hemo-ilmenite (FeO.TiO2-Fe2O3) mineral that is subsequently smelted to produce pig iron and titanium oxide slag. The ore is crushed, heat treated (roasted) and concentrated using dry magnetic separators. This paper describes a model to predict the magnetic susceptibility of the roasted ore and its behaviour in the magnetic separation plant. The model uses a classification of the ore as a function of the size and magnetic susceptibility of the particles. A subsequent division into mineral species allows the simulation of the magnetic separation process. The model is used to assess various operating options when process disturbances occur.INTRODUCTIONOre processing operations consist of ore size reduction followed by the recovery of valuable species in a concentrate that is subsequently sold for its minerals or metal content. The concentration processes are based on the physical and/or chemical properties of the minerals contained in the ore. Flotation is typically used to recover minerals of base metals such as copper, zinc, lead or nickel. Iron ore and other dense minerals are typically concentrated by gravity separation processes. Magnetic separation uses the magnetic properties of the minerals to concentrate magnetite and other minerals, to clean coal concentrate and to recycle ferrosilicon for dense separation processes (Brophy, 1985).Magnetic separators can be divided into low and high intensity separators and high gradient magnetic separators (Gillet, 2008). Most of the low gradient magnetic separators are built as rotary drums that are operated in wet or dry conditions. Dry magnetic drum separators are used by Rio Tinto Fer et Titane in Sorel-Tracy, Canada to concentrate hemo-ilmenite from the ore extracted at the Lac Tio Lake mine near Havre St-Pierre, Canada. The analysis of production data suggests that process improvement could be achieved by identifying key variables for the control of the magnetic separation circuit and/or by modification to the circuit flowsheet. The use of a process model was considered to be an effective approach to conduct a preliminary analysis of the circuit with a view toward process optimization. However, the literature contains few detailed studies dealing with the modelling of magnetic separation processes."

Jan 1, 2011

By Charles Kocsis

Over the last few years a multi-faceted feasibility study has been evaluating the introduction of hydrogen fuel cell powered equipment into underground mines to replaced diesel engine powered equipment. The analysis has shown that under normal operating conditions fuel cells could at the very least be an environmental and health benefit by eliminating combustion engines and their unwanted by-products. However, they could also generate significant reductions in the amount of ventilation a mine needs to supply when compared to dieselized operations; this would further decrease a mine?s energy consumption, its associated greenhouse gas impact and also reduce a mine?s significant ventilation associated costs. Fuel cell powered equipment also has the advantage of lower noise and heat production compared to their equivalent diesel powered counterpart. The degree to which fuel cells can generate ventilation benefits in underground mines depends upon a number of operational parameters and mine specific qualifiers such as current ventilation control or management, mine depth, inherent dust conditions and minimum velocity requirements. However another major consideration in the introduction of fuel cells in underground mines will be the safety requirements associated with diluting and removing the potentially explosive hydrogen in the event of a gas leak. In outside surface applications, it is possible to use dispersion, buoyancy, time, and lack of ignition sources as mitigating factors in deriving a low risk of an explosion. However, in underground operations these factors change, dispersion and buoyancy are limited; time may not be available and in non-coal mines ignition sources are not typically controlled. When a hydrogen leak occurs, the availability of sufficient ventilation to dilute and effectively remove the gas will be critical, consequently ventilation may be the prime risk controlling factor. In this paper, the authors evaluate the potential benefits of replacing diesel engines with fuel cells in powered production equipment, discuss the mitigating qualifiers that could limit ventilation savings and evaluate solutions to retain and maintain an additional ventilation capacity in the event of an emergency situation such as a hydrogen leak from the fuel cell stack or its distribution system.

May 1, 2007

By Hassan Ghaffari

? HPGR Process/Overview ? HPGR vs. SAG ? The Manufacturers ? Mine Operations using HPGR ? Projects by Wardrop/Tetra Tech ? Economic Benefits ? Questions Main Operating Parameters ? Press Force (N/mm2) ? Roll Speed (m/s) ? Moisture Content (%) ? Feed Particle Size Distribution Specific throughput M* = throughput/D*L*V (ts/hm3) Flake thickness/strength in balance with the compressing force (floating roll situation)

Aug 1, 2013

By H. Harzer

Statistics Canada record s indicate that in recent years 70 to 80% of the entire Canadian coal production originated from Western Canadian mines, in which about 55 to 65% of the coals are bituminous. Until recently, less than 20% of the latter were extracted by underground methods , the remainder being surface-mined. However, it is estimated that within the foreseeable future up to 80 or 90% will have to be gained by underground techniques. Although Western Canadian coal resources are considerable, they are unfortunately locked in severely distorted, steeply pitching and faulted seams of variable thickness. Consequently, their economic extraction poses serious engineering problems. One possible solution to the problems is to apply hydromechanical methods to underground mining. This is a well-proven method shown to be economically viable under Western Canadian conditions. This paper endeavours to widen the scope of published data on hydraulic coal mining and is based on practical engineering and cost data gathered through contacts with West German designers and operators of hydraulic mines.

Jan 1, 1978

By J. Dong

"Xanthate is a sulfide collector that has been widely used in flotation for almost a century. It is inexpensive, easy to synthesize and apply, and produces good metallurgy in the right applications. However, a major drawback of xanthate is that it decomposes easily and generates CS2 gas in transportation, storage and operation, which can be harmful to human health. Also, excessive xanthate in tailings pond leads to toxicity of the process water and potential damages to aqua-life systems. New regulations in Ontario require that the TWA (time-weighted-average) value of CS2 in the work place is below 1 ppm, which is 1/10 of the previous limit. The stringent regulation may eventually sweep other regions of Canada with the increased awareness of healthy working conditions and environment protection. Therefore, it becomes very important to find xanthate replacements which will be “green”, i.e., environmentally friendly.In this research work, a systematic procedure is used for collector evaluation using the Sudbury nickel sulfide ores from Vale’s Ontario operations, starting with the collector properties (MSDS) evaluation, followed by batch flotation tests, miniplant tests, plant trials and economic evaluation. It is found that the most promising xanthate replacement is dithiocarbamate.INTRODUCTIONSulfide mineral flotation has been practiced since the early 20th century. Its industrial importance is well recognized as mineral concentrates from flotation can be economically smelted and refined to provide primary metals. Collectors play a vital role in froth flotation. There are reagents possessing both a polar group that binds them to the mineral surface and a nonpolar group that is oriented away from the surface to attach air bubbles. The introduction of xanthate as a sulfide collector in 1925 was a remarkable development in froth flotation (Keller, 1925). It is low cost, easy to synthesize, water soluble, simple to control in application. Most importantly, it has strong capability and usually good selectivity for sulfide mineral flotation, which have been keeping its application prominent for almost a century."

Jan 1, 2011

By C. B. Webber, W. A. Grell

"Sherritt Gordon's Ruttan Mine is a copper-zinc operation located approximately 600 miles north-west of Winnipeg. Open pit mining began in 1973 at 10,000 Tpd., with a change over to underground mining in 1981 producing 7,500 Tpd. Ore is crushed to -3/4"" followed by rod and ball mill grinding (two ball mills in series), differential copper-zinc flotation, and dewatering. The copper and zinc concentrates are trucked 75 miles to Lynn Lake, where they are shipped by rail to the HBM&S smelter at Flin Flon.This paper describes a method used to predict the effect of grind particle size on metallurgical performance, as well as details on the copper and zinc metallurgical performance due to fine grinding in actual plant operation.Initial statistical studies on historical data (including ore source, head gr3de, grind, etc.) indicated that copper and zinc recoveries increased with fineness of grind, in the range of 50 to 65% -200 mesh.In order to extrapolate beyond 65% -200 mesh, and to develop expected nonlinear relations between grind, Cu concentrate grade, and recovery, the assays of screen fractions of copper concentrate and tailing were used to determin8 the relation between net smelter return (a function of grade and recovery) and grinding cost. The optimum economic grind would be that grind at which the net smelter return less grinding cost reached a maximum."

Jan 1, 1982

By G. A. Kosick, G. S. Dobby, O. Matwijenko

"A 100 kg/hr portable flotation pilot plant has been operated at Kidd Creek Mines with the objectives of:(a) determining performance expected by column flotation circuits on primary zinc and copper cleaning, and (b) evaluating column flotation circuit configurations.The self-supporting pilot plant, owned by Minnovex Technologies Inc. and operated by Minnovex and the University of Toronto, consisted of three 16 cm diameter by 6.5 m high columns and a bank of four Denver No. 7 cells. It was configurable to any arrangement using these four (or fewer) units. The columns were controlled through the use of multiple pressure transducer measurements. This paper presents the approach to testing and the results of more than 70 circuit tests. INTRODUCTIONThe integration and retrofitting of flotation columns within existing mechanical cell flotation circuits is becoming a common occurrence. The correct approach to this integration, however, is often difficult to define, due to a scarcity of previous experience and the difficulty in predicting column circuit performance. In most cases, the effect of column implementation has been experimentally assessed using a single pilot column, with the consequence that selection of an appropriate circuit to implement is difficult. Important considerations are the impact that column water addition will have on the existing operations, and the differences between mechanical cell and column performance in integrated cleaning circuits.To address the problem of circuit selection a portable flotation pilot plant, consisting of both columns and mechanical cells, was designed and constructed by Minnovex Technologies. The pilot plant was structured for complete flexibility of circuit configuration, with the intent to provide a means by which both research and plant personnel could determine the best possible column circuit configuration for a given application."

Jan 1, 1990