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By V. Srivastava, T. Ghosh

The modeling and simulation of semiautogenous (SAG) mills have been widely used in the design and optimization of mill performance in terms of its power draw, processing capacity and product size distribution. However, these models are solved under steady approximation and do not provide any information on mill charge distribution in real time. 7KLVSDSHUDWWHPSWVWRFKDUDFWHUL]HPLOOFRQWHQWVE\VROYLQJ:KLWHQ¶V¿UVW RUGHUFRQWHQW EDVHGPRGHOLQWKH0$7/$% Simulink environment. The parameters in the model, such as breakage rate constant, discharge rate and appearance function, were estimated using process and design data collected from a gold mine operating in Alaska coupled with a nonlinear parameter estimation scheme. This model was then used to predict the dynamic responses of other key operational variables, such as mill power, bearing pressure, charge level and product size distribution. The transient response of mill behavior with respect to changes in feed size distribution, tonnage and mill feed water is also presented. This dynamic simulation approach can be used for practicing different control strategy and training purposes. The observed response of mentioned variables was validated using dynamic response data from the plant to encompass operational characteristics of SAG mill behavior.

"INTRODUCTION Twin Metals Minnesota (TMM) (operating as a joint venture between Duluth Metals and Antofagasta PLC) assumed control of the property in 2010. Amec became involved with the project in 2011 and in 2012 incorporated additional drilling in order to update historic resource estimates. The Maturi geological model and resource estimate was updated in 2014 to incorporate additional information obtained during 2013 and 2014 drill campaigns. In 2017, Amec was purchased by the Wood Group and is now known as Wood. Geology The Twin Metals Minnesota (TMM) Maturi, Maturi Southwest, Birch Lake, and Spruce Road properties are located in northeastern Minnesota (Figure 1). This talk discussed the Maturi deposit only, although all four deposits share similarities. The following geology is summarized from Peterson and Severson (2002).Maturi lies within the Superior Province of the Canadian Shield which is exposed in central and northeastern Minnesota and extends northward into Ontario. In eastern Minnesota, the Midcontinent Rift System developed during crustal scale extension in the Mesoproterozoic Era. The rift system is traceable, as exposures of mantle-derived tholeiitic to subalkaline mafic lava flows, intrusives, and rift-filling fluvial sedimentary rocks, and in the subsurface as a gravity anomaly (high), from the eastern end of Lake Superior, arcing west across the lake basin, and extending south-southwest to northeastern Kansas. Intrusion of the Duluth Complex, circa 1,108 - 1,099 Ma, was related to rifting and is co-genetic with the North Shore Volcanic Group mafic volcanic rocks that form the hanging wall of the Duluth Complex to the east. Cu-Ni-precious metals (PM) mineralization including Pt, Pd, Au and Ag within the Duluth Complex is hosted by mafic to ultramafic intrusive rocks that are part of the overall intrusive complex. The Duluth Complex is defined as a more or less continuous mass of mafic to felsic plutonic rocks that extends northeastward for more than 170 miles (275 km) in an arcuate fashion from Duluth nearly to Grand Portage in Minnesota. It is bounded by a footwall of Paleoproterozoic sedimentary rocks and Archean granite-greenstone terranes (Peterson and Severson, 2002), and a hanging wall largely of comagmatic anorthosite, rift related flood basalts, and hypabyssal intrusions of the Beaver Bay Complex. Mineralization is hosted in the basal mineralized zone (BMZ) of the South Kawishiwi Intrusion (SKI). The SKI magmas intruded sub-horizontally between hangingwall Anorthositic Series (AN) rocks and footwall granitic rocks of the Neoarchean Giants Range batholith (GRB). The Maturi local geology and the deposit footprint (shaded orange) is shown in Figure 2."

Jan 1, 2018

By C. E. Jordan

Direct digestion of phosphate ores by the acid-alcohol method offers improved leaching selectivity for the exclusion of iron and aluminum in the product acid. The leaching kinetics of this process will have a major role in the eventual application of this technology. The Bureau of Mines investigated apatite dissolution to determine the reaction rate and mechanisms of the leaching kinetics. The dissolution kinetics were monitored on various sized fractions of pure apatite. The "shrinking-core" kinetics model was applied to the data to determine the rate controlling mechanisms. For each size fraction, the initial reaction rate was controlled by the dissolution reaction. After an hour the dissolution rate slowed indicating a diffusion controlled reaction caused by the formation of calcium sulfate hemi-hydrate coating on each particle. For complete dissolution of the apatite, the phosphate rock should be ground to near 10 µ m size.

Jan 1, 1993

By J. Marbeau

In this paper, a geostatistical study of the North Butte roll type uranium deposit is presented. The geostatistical procedure applied takes into account the varying direction of the roll front for reserves calculations. For variogram calculation, the roll front is unfolded thereby allowing estimation of the variogram in directions parallel to and perpendicular to the roll front. Variograms calculated in unfolded space show strong directional anisotropies consistent with the anticipated configuration of the deposit.

Jan 1, 1991

By I. H. Gundiler

Although cyanidation has been universally applied in the processing of gold and silver ores, the search for less toxic alternatives continues. Difficulties encountered in processing ores containing oxidized copper minerals and the inherently slow kinetics of cyanidation have generated a renewed interest in thiosulfate leaching. This process, as well as the catalytic effects of copper ion, were well recognized and widely applied towards the end of the past century. In this study, the effects of copper, ammonia, and thiosulfate concentrations on gold dissolution were investigated using the rotating disk technique. Extraction of gold and silver from an oxidized copper-bearing ore was also studied utilizing a statistical experimental design, and these results were displayed in the form of contour plots.

Jan 1, 1993

Current scrubber capacities vary from 1.89 to 4.72 m3/s (4,000 to 10,000 cfm), and there is interest in using scrubbers with capacities up to 6.61 m3/s (14,000 cfm). This paper describes research to compare the effectiveness of blowing and exhausting face ventilation systems using high-capacity scrubbers with flows up to 6.61 m3/s. Tests were conducted in a surface ventilation gallery with exhaust and blowing systems operating under similar conditions. The mining machine was equipped with a machine-mounted scrubber. Methane gas was released at constant flow at the mining face and measured at several face locations. Effectiveness of exhausting and blowing systems was compared by determining average face methane concentrations. For each operating condition tested, blowing ventilation was more effective than exhausting for controlling face methane levels. Raising scrubber flow to 6.61 m3/s reduced methane levels for both blowing and exhausting systems.

Jan 1, 1996

By Z. Man, X. Zhu, Y. Xian, Y. Tao

"Fine coal separation has been always concerned. The froth flotation is not efficient in desulfurization because of the low rate of sulfide liberation from raw coal, the natural hydrophobicity caused by sulfur concentration on the surface of sulfide, and the presence of large amounts of middling feed. An efficient centrifugal gravity separator was employed to separate for Shandong –0.5mm high sulphur coal, the partition curve of centrifugal separator was obtained with conclusion that the Ecart probable error value of the separator is 0.095, the imperfection I value is 0.128. Finally, comparing with the coal flotation using progressive release method, the result shows that the effect of the desulfurization with centrifugal gravity force field based on density difference between coal and minerals are much better than that of the floatation which operates based on the difference in surface properties. It provides a efficient technical way for fine coal separation. INTRODUCTION Coal plays the most important role in the energy consumption in China. Although natural gas, wind energy, nuclear power has made great progress in recent years, but the share of coal in the final energy consumption accounts for 70-80%, which is determined by energy storage of China. Under the requirement of environmental protection, the clean utilization of coal resources becomes more and more important, and the need has already been identified. Coal washing and processing plays an important role in the process of high-efficiency and clean utilization of coal. For the separation of fine coal, flotation is a traditional method for effective separation and has been used in large scale industrial applications in coal preparation plant. However, for the separation of difficult-to float coal, especially fine oxidized coal and low–rank coal with hydrophilic surface, flotation is usually inefficient [1-3]. Although many pretreatment methods have been adopted, including grinding, microwave pretreatment, new flotation reagents, the problem of efficient separation has not been completely resolved [4-6]. Thus, an enhanced gravity separation method using particle density difference rather than hydrophobic difference was proposed for the separation of fine coal. Many research results have proved that this is a practical and effective method of sorting. R.Q. Honaker used Falcon C10 concentrator, Knelson centrifugal separator and spiral separator to clean the fine coal with size of 600*210 µm and 210*37µm, the effects of different operating parameters on the separation index was discussed. The results show that the separation density of Falcon can reach 1.6, and the Ep value can reach 0.1-0.15. This paper revealed that for the separation of fine coal with high sulfur content or fine grained content. Falcon has high separation efficiency with low cost. Separation results present in another paper showed that the desulfurization efficiency of inorganic sulfur was better than that of flotation [7-8]."

Jan 1, 2017

By S. Tafazoli, D. Cooper, I. Bell, S. Nouranian, B. Sameti, M. Ramezani

"Image-based rock fragmentation sensing in mines and quarries includes an important rock boundary delineation step, commonly referred to as rock segmentation. This paper presents an artificial intelligence-based solution to this challenge. The proposed technique encodes prior knowledge of previously analyzed images into mathematical/statistical models. Human-labeled images are used as inputs to train neural networks through an optimization process. The networks can then be used in real time for rock delineation. To build the models, special deep artificial neural networks are used as a pixel classifier. The proposed classifier labels each pixel (edge, rock or fine) by analyzing a plurality of pixels within the image. Advances in machine learning allow the network to contain many parameters. The increased number of parameters is a strong factor in the classifier’s ability to correctly classify each pixel. Deep learning-based segmentation, combined with 3D imaging followed by post processing, provides a unified fragmentation sensing solution. Results from automatic segmentation are compared to human labeled segmentation using the percentage passing curves for 64 rock images. INTRODUCTION Mines and quarries can see significant productivity and performance benefits by controlling material sizes (McKee, Chitmobo, & Morrell, 1995; Sellers & Gumede, 2012). During the past decade, image-based fragmentation analysis has been applied to estimate rock size distributions in order to optimize procedures in mines and quarries. Using a roving camera and operator assisted analysis Maerz, Franklin, and Coursen (1987) measured the size distribution of the blasted rock and eliminated the need for manual sieve analysis. Since then, many others have introduced automated and manual methods for fragmentation analysis and improved upon existing approaches (Girdner, Kemeny, Srikant, & McGill, 1996; Smith & Kemeny, 1993; Palangio, Palangio, & Maerz, 2005; Tafazoli & Ziraknejad, 2009; Raina, 2013). In image-based fragmentation, rock boundaries are identified in the image, and an image scaling is applied to transform rock pixel sizes into real world dimensions. Usually geometric references, such as regularly shaped objects (discs and basketballs) (Siddigui, Ali Shah, & Behan, 2009), or the known size of an excavator bucket (Zeng, Chow, Baumann, & Tafazoli, 2012), are used to determine the proper scaling factor. In recent years, 3D imaging and sensing have also been incorporated to improve rock delineation. Specific methods include using camera-laser combinations and stereo imaging to measure rock fragmentation on conveyor belts (Noy, 2013; Thurley, 2013; Dislaire, Illing, Laurent, & Pirard, 2013) as well as portable fragmentation analysis devices with 3D imaging sensors (Sameti, et al., 2014)."

Jan 1, 2018

By C. Haycocks

The evolution of longwall mining is briefly traced from its inception in England in the 1620's. In the beginning this method recovered 90% of coal with roof closure controlled by packs and walls. Mechanization began with kerf cutting, crude face conveyors, and improved face support. Full face mining was accompanied by the introduction of self-advancing hydraulic supports. U.S. longwalls evolved a retreating system to recover blocks of coal outlined by room and pillar panels. Computer control combined with the most modern technology has greatly increased productivity. Past trends are used to predict future technological directions.

Jan 1, 1997

By Ö. Y. Gulsoy, L. Ergün, I. B. Celik

"In this study, the studies to develop a flowsheet for a low grade chromite ore are presented. The existing concentrator had to be suspended due to the decreasing metal prices. The study commenced with mineralogy and liberation profile of the chromite. The valuable mineral is chromite and the major gangue mineral is forsterite together with small amount of brucite and calcite. Heavy liquid tests were performed on different size fractions without any acceptable success. The concentration tests were performed using teetered bed separator (TBS) +shaking table, spiral concentrator+TBS+shaking table for different grind size. Based on the test results, 100% finer than 0.4mm was found to be optimum for grind size. Comparison of alternative circuits showed that the performance of both alternatives would be similar. A flowsheet was designed and operated for as 25 tph trial circuit in the existing plant. The results showed that the performance of the plant predicted was very close to the actual plant data. The paper describes the methodology used, presents the experimental data and actual performance figures.INTRODUCTION Turkey is one of the largest chromite producers in the world. Since 1850, chromite mining found industrial applications and geographically spread along six different regions in Turkey. Although the reserves in these locations are not as rich as the big producers, they are regarded as very valuable due to Cr/Fe ratio in terms of chromite utilization standards (1, 2). In this period, significant portion of high grade (=35% Cr2O3), directly marketable ore deposits have already been produced. According to General Directorate of Mineral Research and Exploration (Turkey) data; 83% of the known chromite deposits of Turkey which makes 242 341 000 tons in total, is consisting from the deposits those have a Cr2O3 grade lower than 10%. Only 5% of the known deposits (13 087 000 tons) has a grade of directly saleable (2, 3). Gravity concentration methods are well-known processes for chromite beneficiation and there are plenty of plants in the world and Turkey using these methods. As different equipment combinations can be used for the same enrichment method, different methods can also be used in combination for plant designs (4). Beneficiation of low grade and problematic chromite ores may lead to use these method and equipment combinations (5, 6). Ore liberation, the grades of Al2O3, SiO2, FeO and Cr2O3 in the obtained concentrate, Cr/Fe ratio, some physical and chemical properties between gangue and minerals (i.e. specific gravity, magnetic property, color, etc.) designate the selection and combination of these methods and equipment (7)."

Jan 1, 2017

By Cuthbert Musingwini, Adeodatus S. Nhleko, Mahlomola I. Mabala

Optimization of a three-dimensional (3D) underground stope layout is complex given the number of variables and constraints to consider, such as the stope size. Determination of the appropriate stope size in stope layout optimization problems is a complex process that must consider several factors including placement of infrastructure and development, production scheduling, and equipment selection. An inappropriate stope layout can result in unrealistic expectations for a production schedule, consequently negatively affecting the derived economic value. Therefore, this paper used a dual interchange algorithm (DIA) to evaluate the effect of stope size on the stope layout economic value (SLEV) for a synthetic Platreef mineral deposit planned for extraction using sublevel open-stoping. The results showed that generally as the stope size increases, the SLEV decreases. This aligns with practical mining because larger stope sizes have a higher chance of incurring more dilution which reduces economic value, while smaller stope sizes enable more flexible extraction where unprofitable blocks can be left unmined as waste, thus increasing the derived SLEV.

Jun 25, 2023

By Luis R. Bittencourt

Research efforts have been made to characterize a refractory Brazilian bauxite ore composed basically of gibbsite, kaolinite and quartz. Detailed polished-section microprobe analysis of the unbroken ore has been used to describe the liberation characteristics and the association between mineral phases. The linear grade distributions for selected particle size fractions were determined by means? of a particular implementation of King's Random Fracture Model that uses the exponential model for the distribution of linear intercepts through both dispersed (kaolinite) and continuous (gibbsite) phases. The corresponding volumetric grade distributions were then determined by means of a constrained optimization technique for the solution of the transformation equation. Batch flotation experiments were carried out to determine the kinetics for particles of different sizes and the results correlated with the liberation spectrum for each size class.

Jan 1, 1991

By Jon Bowler

Daw Mill Colliery, UK Coal, near Coventry in England, mines the Warwickshire Thick Seam at 800 m/2500 ft depth using longwall retreat with gateroads supported by rockbolts. Over the last eight years the mine has worked longer and wider panels with extraction heights of over 5 m. Recently the mine has experienced high deformation conditions principally from the rib sides resulting in operational issues and production delays. Investigation of rib deformation was undertaken using new inspection and measurement techniques, including systematic laser surveying, and improved rib support consumables were developed and introduced. A new rib trussing system using textile webbing has proved very successful.. A rib support management system was also developed and introduced for face retreat. This allowed rib support to be more effectively managed both for production and safety. For the current developments, new 28-mm-diameter roof and rib bolt systems have been introduced and rib bolt length has been increased to 3 m, with the aim of achieving gateroad stability over the life of the panel. Significant reductions in roof and rib deformation have been confirmed by the latest monitoring results. These ground control innovations are allowing Daw Mill?s potential to be realised. Production from 302?s face in 2008 set new UK and European records. However the support enhancement process continues and further innovations are planned.

Jan 1, 2009

By M. I. Brittan

A technique has been developed that analyzes routine mill operating data to determine the key operating parameters responsible for gold losses to tails. These parameters represent the prime operating variables having a dominant effect on gold recovery. The technique quantifies these effects for the specific plant being analyzed, providing a tool to optimize recovery and costs and so enhance the mine?s cash flow. The quantitative relationships generated will be different for each ore and plant: every case is site-specific and has to be treated on its merits. Examples are provided to illustrate the considerable metallurgical and financial benefits which can be realized from optimization. The examples are taken from case studies of mills in four major gold mining districts of the world.

Jan 1, 2010

The knowledge of aggregate size under various chemical and hydrodynamic conditions is very important for selective agglomeration, flotation and water clarification processes. Traditionally, aggregate size measurements are conducted using an ex-situ technique or in a diluted suspension. In either case, the measurement occurs in an environment that does not resemble the actual conditions under which the aggregates are formed. In this study, an on-line scanning laser microscope has been used to study the coagula size of micronized coal suspensions as a function of pH, solids concentration and agitation intensity. A maximum aggregate size of 310 µ m was obtained under relatively quiescent conditions and at pH values near the iso-electric point (iep) of the coal. The aggregate size results compare well with the coagulation efficiency measurements obtained under the same pH conditions.

Jan 1, 1994

By Fun-Den Wang

In underground coal mines, three types of roof failure (2)*, tensile, local, and shear, occur in the development entries and in room-and-pillar extraction areas. Most of the tensile failures in the roof at the center of the mine opening as well as those local failures due to fracture or jointing can be controlled by roof bolts, trusses, and timbering. However, control of the shear failure at the sides of the roof (or the cutter-roof caving) has not been successful with ordinary roof control methods. The only effective method, when there is an indication of cutter-roof caving, is to use cribs, which are costly and make transportation, ventilation, and production difficult. Frequently, the operator allows the cutter roof to cave and then cleans it up rather than trying to stop it. This shear type of roof failure is often associated with large areas which may progressively involve several hundred feet of roof caving at one occurrence. At times, this shear failure occurs suddenly with little or no warning. This can result in buried equipment and great danger to the safety of personnel. This shear failure may be due to one or more of the following: high stress concentrations at the corner of the roof from (1) high horizontal to vertical load ratio, (2) orthotropic elastic properties of rock and coal, (3) greater stiffness of the immediate roof layer than of the secondary roof layer, (4) viscoelastic behavior of rock and coal, and (5) geometry and material properties of the geologic setting. Removing or reducing the high stress concentration should improve the roof stability. The structural analyses of models of two typical

Jan 1, 1972

By S. Anderson

Advances in inexpensive computer automation led to the development and utilization of a computerized ore control system at Cyprus Coppers tone Gold Mine near Parker, Arizona. These new ore control practices improved accuracy, performance, and productivity of the ore control effort. An automated system for data collection, including bar coding of blast hole samples, laboratory analysis, data management, grade estimation, cut delineation, and ore zone stakeout, plays a major role in ore control practices. This is of particular interest because of the large amount of blast hole information being processed, analyzed, and reported daily.

Jan 1, 1991

The Kinross DeLamar Mine has been operated for more than 17 years producing precious metals dore' from ores, principally silver bearing with associated gold co-products. In late 1995 a new orebody was commissioned, the Stone Cabin orebody, and the site began a phased shift from silver dominance to gold dominance in terms of precious metals revenue. The Knelson Concentrator was selected to remove gold from a portion of the cyclone underflow stream in the grinding circuit. This paper describes the development of the flow sheet, specific mineralogical impacts on Knelson product and downstream materials, and impacts on mine/mill grade reconciliation.

Jan 1, 1997

By N. D. Jagger, I. M. Arbuthnot

Introduction Over the last five years, a large number of small- to medium-sized carbon-in-pulp treatment plants have been built in Australia, most designed to treat between 250,000 t/a and 1.5 Mt/a of ore. Because of the limited capital resources and tight cash-flow positions of these relatively small mining companies, the primary requirement was often to get a plant built and operating in a short period of time and at minimal capital cost. Therefore, since the inclusion of both pre-leach and tailings thickeners represents an obvious and significant capital cost, most of these plants were built without thickeners or even detailed, cost-benefit analyses on their inclusion. In some cases, the increasing use of High Rate Thickeners (HRTs) in the mineral processing industries has, however, resulted in a reassessment, because of their considerably lower cost. This reassessment was triggered primarily by the need to conserve water in arid mining areas where borefields are costly to install and water is limited. With the startup and operation of these installations, the resulting significant savings in cyanide consumption has been recognized, in many situations, as a primary justification for the installation of HRTs. Solution balances Degradation of cyanide occurs in the tailings water discharge to slime dams. The degree of degradation (cyanide loss) in the water recovered depends on a number of factors, but it is usually assumed to be about 90%. The most important mechanisms of CN loss are through HCN losses and oxidation by oxygen in the air, which also assists in the hydrolysis of CN. These mechanisms are supported by the large dam surface area and the long retention time of the tailings water in the dam. By thickening the CIP tailings at the plant and recovering as much tailings water as immediately possible, these losses are avoided. The retention time in an HRT is less than three hours, and the surface area is relatively small. Therefore, CN losses are negligible, which is not necessarily the case in conventionally-sized thickeners. Fig. I shows a block diagram of a 100-t/h gold plant without a thickener. In this example, 50% of the tailings water pumped to the tailings dam is recovered, and the CN concentration of the returned water is 10% of the tailings CN concentration of 150 ppm. Fig. 2 represents a plant with an HRT on tailings, thickening to 55% w/w solids. In this case, the thickened tailings are pumped to the dam, and 25% of the contained water is recovered. The recovered water and the mill make-up water are not sent directly to the mill; instead, they are added to the thickener feed and mixed with it prior to thickening. By doing this, the tailings are effectively washed, and the additional cyanide is recovered. Solution balances over these two circuits show cyanide recoveries of 5% and 65%, respectively. Thus, the thickener use increases cyanide recovery by 60%. Fig. 3 shows a two-stage HRT circuit in a countercurrent decantation (CCD) configuration. In this example, an additional 13% of cyanide is recovered through the use of the second-stage unit. This configuration can be justified when residual cyanide levels are high. Capital and operating costs - Case study 1 Illustrative cost figures are based on a CIP tailings thickener installed, in early 1988, as part of Dominion Mining's treatment plant at Paddy's Flat near Meekatharra. (All costs within this paper, unless stated otherwise, are in Australian dollars.) [Assumptions: Feed rate 150 t/h Ore moisture content 5% Leach density 40% solids Underflow density 55% solids Residual cyanide in tailings 150 g/m3 Flocculant dosage 15 g/t Consumable costs: Water $ 0.40/m3 Cyanide$ 2.00/kg Flocculant$ 4.50/kg Power$ 0.12/kWh Length of tailings pipeline2,500 m Capital costs of the major items involved in the thickener installation are given below: Thickener unit, 15-m diam $ 230,000 Water return pumps 12,000 Water recirculation pumps 28,000 Feed box 5,000 Flocculant make-up system 18,000 Flocculant storage tank and dosing pumps 11,000 Piping and valves 85,000 Electrics and instruments 47,000 Civil works 29,000 Installation 69,000 Total cost$ 534,000] Savings in capital costs that can be attributed to the

Jan 1, 1993

By Andrew N. Peterson

US mine operators frequently fail to investigate more cost effective and productive bulk material handling systems because surface mines seem to lend themselves to truck ore haulage. In this country, as a result, use of conveyors to move heavy loads from mine to process facilities has been minimized, if not actually neglected. In contrast, there are more than 50 conveyorized surface mines in successful operation around the world. These mine operators have learned that properly applied conveyorized systems can offer major savings in capital and operating costs, which contribute to improved profits when combined with other proven mining technologies. Growing acceptance and application of conveyorized bulk material handling in surface mines also points up how unique each mine is and how careful planning contributes to maximum mine effectiveness. Because of these differences, mining executives and technical and operating staffs need to develop an understanding of three factors in applying conveyorized bulk material handling in surface mines: • Why each mine will benefit from the type of automation permitted by conveyorized operation, •What kind of equipment is available, and • What applications most effectively demonstrate the first two factors in action - hauling either ore or waste. The conveyorized systems considered in this presentation have production rates from 0.5-2.7 kt/h (500-3,000 stph). Worldwide, these systems have been operating since the early 1960s. Advantages of Conveyors Why do you want conveyorized bulk material handling? First, it almost always provides lower operating and maintenance costs. Second, it frequently requires lower initial capital costs and almost always requires lower capital costs over the life of the surface mine. Third, it provides comparable operating availability, and finally, it frequently gives comparable operating flexibility - depending on the mine plan. Cost avoidance can be accomplished with modern production methods. These, in turn, permit increased productivity and reduced operating costs such as those for energy, maintenance, and manpower. It has been demonstrated in European surface mines and elsewhere, that conveyor systems frequently require lower initial costs than does truck haulage. Almost always such operations require lower capital costs over the mine life. Those costs include the continual addition of haulage trucks to both accommodate the increasingly difficult haulage routes and fulfill replacement requirements when trucks wear out. Conveyor systems handling ore in numerous large crushing and port facilities, which have operated since the early 1950s, have clearly demonstrated a useful conveyor life of more than 25 years. In contrast, off-highway trucks have life spans of six to eight years. The following examples illustrate comparative capital costs to purchase conveyor systems and comparable truck haulage units. Example 1 The ore haulage route from point A to point B is level and 610m (2,000 ft) long. The material weighs 1.8 t/m3 (110 lbs per cu ft) and must be transported at a rate of 1.8 kt/h (2,000 stph). The installed capital costs to provide a properly designed conveyor that will transport the described material from point A to B is about $450,000. The capital cost to purchase three 77-t (85-st) off-highway trucks and one spare truck - which would provide equivalent capacity - would be about $1.2 million. The truck cost estimate is based on a 6 min. or 771 kt/h (850 stph) truck cycle time. Truck efficiency is estimated at 0.8. Each 77-t (85-st) truck would have an actual haulage rate of 617 kt/h (680 stph). Therefore, three trucks would be necessary to transport the designated tonnage of 1.8 kt/h (2,000 stph). A movable crushing plant would be located at point A for the conveyors and a permanent crushing plant at point B for the truck haulage system. Capital costs for these primary crushing plants were not included in the calculations for either system because the capital costs are frequently comparable. Example 2 The transport route from point A to point B is 610 m (2,000 ft) horizontally and 122 m (400 ft) vertically - on a 20% grade (Fig. 1). The material weighs 1.8 t/m3 (110 lbs per cu ft) and must be moved at a rate of 1.8 kt/h (2,000 stph).

Jan 6, 1983