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By Juan J. Monsalve, AMAN SONI, ALEX PFREUNDSCHUH, Nino Ripepi

Ground control failures are one of the main causes of accidents in the underground stone mining industry. Some of the fundamental tools for rockfall hazard identification are related to rock mass characterization and geotechnical discontinuity mapping. Recent technological advances in these methods are related to remote sensing techniques and pointcloud processing software for automated discontinuity mapping. Remote sensing techniques, such as LiDAR and photogrammetry, generate multimillion-point clouds with millimetric precision, capturing the structure of the rock mass. The automated point-cloud processing tools offer alternative algorithm-based methods to characterize and map these discontinuities. However, their applicability is constrained by multiple factors, such as site-specific conditions of the rock mass and the parameters used within the mapping algorithms. This paper evaluates the performance of automated discontinuity extraction software compared with manual virtual discontinuity mapping. Sampling windows from laser-scanned sections in an underground limestone mine are defined and mapped using the Discontinuity Set Extractor (DSE) opensource software. Results from the virtual discontinuity software are compared with manually extracted fractures from I-Site based on reviewing orientation, trace length, spacing, number of extracted discontinuities, and processing time.

By N. C. Miller

In situ stope leaching is an innovative mining system that reduces the surface impacts of conventional underground mining. As in any leaching operation, stope leaching requires hydrologic site characterization and control of leaching solutions. Several tracer tests, using sodium chloride (NaCl) in various concentrations, were conducted by the U. S. Bureau of Mines (USBM) in an underground experimental stope to simulate a leaching environment. The tests were conducted in fractured crystalline rock that was partially saturated. As tests progressed, the surrounding fractures became more saturated. A predominant flow direction away from the stope was found during testing. Testing procedures, preliminary results, and remediation ideas based on field observations are presented in this paper.

Jan 1, 1995

By Joseph W. Murray

GENERAL DESCRIPTION The undercut-and-fill method of mining was devel¬oped by Inco Ltd. in the Sudbury district of Ontario, Canada, to deal with abnormal ground conditions en¬countered in pillar recovery. Although it is now some¬times utilized as a primary stoping method, the main application continues to be in pillar recovery. In any mining method in which fill is employed, pillar systems tend to fail gradually as stoping pro¬gresses. As a result, most pillars are completely destressed and broken by the time pillar recovery com¬mences. Control of the back in such ground becomes a major problem and is contingent on local conditions as they exist in the workplace. In most mining methods, operations such as drilling, blasting, slushing, and tim¬bering can be readily assessed and the mining cycle pre¬dicted. However, in the mining of pillars with overhand methods, the time and effort required to maintain a safe back varies to a quite unpredictable extent. For years the square-set method was the most effective means available for pillar recovery. However, in areas where bad ground was encountered, square setting was a costly and sometimes hazardous method. As the tonnage mined from pillars increased, it was apparent that a safer and less expensive approach was required. Past experience indicated that the best way to deal with broken ground was to mine from the top down. Top slicing systems were examined but were rejected due to the major sub¬sidence which resulted. However, ideas generated from studying top slicing finally developed into the present method of undercut-and-fill. This method, which has almost completely replaced square-set mining, exhibits reduced costs, has improved safety and efficiency, elimi¬nates the factor of bad ground overhead, and allows a closer prediction of the actual mining cycle. The result is more accurate scheduling of pillar production. This factor is of increasing importance as the tonnage from pillar recovery increases in relation to total mine pro¬duction. Undercut-and-fill is a method of extracting a block of ore by mining successive cuts, working from the top down. After a cut of ore is completely mined out, lami¬nated timber stringers are constructed along the sides for the full length of the cut. Round logs covered by a wire screen are laid across the stringers to form a mat. The cut is then tightly filled with hydraulically placed cemented sand fill. Mining is then resumed on the cut immediately below the mat. As the cut advances, the timber stringers are supported by round timber posts seated on the solid bottom of the cut. Drilling, blasting, muck removal, and timbering operations are repeated until all the ore in the cut has been removed. At this time, another log mat is laid and the cut is filled. This sequence is continued until the block has been mined from level to level. The basic principles of undercut-and-fill can be adapted to a variety of mining situations. It can be applied transversely or longitudinally in ore bodies of varying width, that dip from vertical to flat lying. The method can be applied using conventional equipment or it can be mechanized through the introduction of small diesel or air-powered units for muck removal. Undercut¬and-fill is a selective method, ore recovery is high, and dilution can be controlled to an acceptable level. How¬ever, it is a high-cost, labor-intensive method and as such its use is restricted to areas where the less costly meth¬ods of ground support are not successful or where bulk mining methods are not adaptable. One basic require¬ment for undercut-and-fill is an adequate source of hydraulically placed cemented sand fill and a dependable sand plant and distribution system. The success of the undercut method is to a large degree dependent on the competence of the fill placed in each of the successive cuts. In theory, the weight of the fill on each cut is supported in part by the cemented sand itself as it binds to the rough sides of the mined¬out areas next to the pillar walls, and in part by the tim¬ber support. As mining progresses beneath the mat, a downward pressure is exerted within the cemented fill and a voussoir arch develops above the mat, across the full width of the cut. This supports the major portion of the load. The log mat and posts have a double func¬tion in that they carry the remaining weight not sup¬ported by the voussoir arch and in addition prevent any pronounced settling of the fill mass. Excessive settling breaks the bond between the cemented fill and the pillar walls, eliminating the arch effect and transferring the entire load to the mat and posts. The theory of a voussoir arch forming within the cemented sand fill mass is substantiated in practice by the fact that pillars have been mined from level to level with no more significant weight on the bottom cuts than was experienced on the top cuts. In effect, as mining progresses downward, there is no additional transfer of weight from previous cuts to the cut being mined. APPLICATION OF UNDERCUT-AND-FILL Because of their size, the Sudbury ore bodies are, for the most part, mined transversely and, to a considerable extent, by filling methods. Transverse stoping blocks are designed with 9-m (30-ft) crown pillars established be¬low each level and with stopes separated by vertical rib pillars. In past years, floor pillars were left above the level as well, but present practice favors the silling of primary stopes at the base of the rail. Rib pillars are designed to meet the ground control requirements of the primary stoping methods. Local conditions and past experience generally dictate their size. Historically, since the square-set method was to be used for recovery, rib pillars were designed to accommo¬date standard 1.7-m (5'h-ft) sets. A typical rib pillar is four sets wide [6.7 m (22 ft)] when located between two square-set or cut-and-fill stopes having a width of 8.4 m (27.5 ft). When the undercut-and-fill method was de¬veloped, a mat width of two sets was adopted. This di¬vided each cut of the common four-set pillar into two

Jan 1, 1982

By M. Ishaq, J. Rosami, J. Brune, E. Sidrow, C. Tsai, S. Slouka

Respirable rock dust poses serious long-term health effects to workers in mining and tunneling underground hard rock environments. When inhaled, respirable silica particles commonly found in quartz and other minerals will scar sensitive lung tissue and cause irreversible lung diseases. Characteristics such as concentration, type of mineral, particle size, and particle shape inhaled can harm workers to various extents. Therefore, this study characterizes rock dust particles less than 100µm in aerodynamic diameter that are released from cutting hard rock in the underground mining environment. In this study, laboratory full scale cutting tests on samples of potash rock were performed with radial picks to generate dust. Three stages of pick wear were tested: new, moderately worn, and severely worn. Comparisons between different stages of pick wear to dust concentration, size distribution, and particle shape characteristics are drawn from this study to start determining the scenarios when the released airborne dusts become potentially most hazardous for workers. Although further testing needs to be conducted to make viable conclusions, the moderately worn pick, or the pick with the most blunt tip, in these experiments are potentially the most dangerous. The blunt pick tip released the highest concentration of dust and generated particle shapes with the highest aspect ratio compared to the other two picks. The preliminary results of this study allow for the further development of optimum bit management and maintenance systems to control the exposure of respirable dust to mine workers.

Mar 2, 2022

By John Dodd

Corrosion studies in laboratory ball mills and grinding environment simulations produce corrosion rates and total metal removal rates much lower than those recorded in operating production ball mills. The corrosion component of ball metal loss in commercial mills represents less than 10% of the total loss expressed as micrometers/hour of exposure in recent USBM tests. Such small absolute losses are consistent with ball metal loss measured in production ball mill marked ball wear tests to determine effects of corrosion and abrasion resisting properties of alloys in operating mills which indicate that metal hardness and microstructure determine relative metal loss rates of a given situation. The combination of ore characteristics and grinding process dynamics determine metal loss severity. Alloying for corrosion resistance and modification of the milling corrosion environment yield relatively small metal loss reductions. Alloying and heat treatment to produce high hardness and microstructures of high carbon martensite and for cast irons, M7C3 carbides produces the lowest metal losses in typical wet grinding applications.

Jan 1, 1984

By D. O. Miller

The effective staffing of personnel in a new concentrator facility is a key factor in the success of the operation. By selectively recruiting quality individuals, a company is in a strong position to build a positive work ethic which will become an integral part of it's culture. And culture, which is a set of standards that all employees work to uphold, is that intangible as- set that often determines whether or not a company will survive during difficult economic times. Recruiting and placement of personnel in a concentrator facility requires an emphasis on planning. Management must carefully analyze the following elements when designing a hiring plan: 1. Number of positions 2. Hiring schedule (timetable) 3. Critical positions 4. Lead-time for trainee posit ions 5. Characteristics of the ideal workforce This presentation will focus on a proven system that is efficient (and yet simple) for achieving a new concentrator facility's manpower requirements. IDENTIFYING THE STAFFING NEEDS As previously stated, the key to staffing is planning. Management will find that a basic manpower chart will provide the hiring schedule to meet deadlines. In fact, this system begins with a manpower chart (Figure 1). The chart basically works in reverse of the hiring requirements: the last scheduled hire is the first job on the list. This approach ensures that the company will not find itself compressed against the facility start-up date. A cardinal rule in recruiting a workforce is to allow sufficient time to orient and train employees to the procedures and culture of the system. Problems in selection of personnel begin when the hiring program is short-circuited for timing reasons. The quality of employee selected invariably suffers accordingly. The second section is the key to the success of the manpower program. This is the latest date by which a formal recruiting effort must begin in order for the hire date to be met. Leadtimes are usually as follows: 1. 90 days for company transfer positions 2. 60 days for experienced new hire positions 3. 30 days for trainee positions

Jan 1, 1986

By E. F. Draskovich

In 1984 a high temperature heat recuperation system was installed in the pellet plant of the National Steel Pellet Company located in Keewatin, Minnesota. The system is similar to those of Cleveland Cliffs Empire and U.S. Steel's Minntac operations. It recovers hot process air from the cooler by induced flow, and directs it to an intermediate zone between the drying and preheat zones on the grate machine. This system recently underwent further modifications in April 1985, following initial operation in 1984. The M. A. Hanna Company, which manages the property, developed the design criteria and made the subsequent recommendations for modification through its research facility in Nashwauk, Minnesota.

Jan 1, 1986

By R. Mensah-Biney

The kinetics of loading of gold from bromide solution onto an anion exchange resin, PAZ-4, were investigated in a batch stirred reactor. The initial loading rate was described mathematically by a first-order reaction rate equation. The effect of agitation speed, temperature, resin particle size, resin concentration, initial concentration of gold and bromine in solution on the rate was investigated. The model described fairly accurately the loading process in terms of the variables investigated. The rate constant for the initial loading ranged from 0.5 to 1.3 x10-3 sec-1 and the half-time was 9.3 to 56 minutes. The initial rate constant increased with increasing temperature with an activation energy of 15.7 kJ/mole (3.8 kcal/mole). The rate constant was directly proportional to the concentration of the resin in solution. The subsequent reduction reaction was also modeled by the first-order rate equation and the apparent rate constant was estimated to be 3.0 x 10-6 sec-1 with a half-time of about 44 hours.

Jan 1, 1994

By Atta Ur Rehman, Kwame Awuah-Offei

This paper examines the effect of bucket geometry (cutting-blade profile, floor profile and blade thickness), speed, motor power output and muckpile particle sizes on bucket penetration and draft for rubber tire loaders. The work uses a 1:16 scale model of a load-haul-dump (LHD) in full-factorial experiments to examine the effect of the variables on penetration and maximum draft. The results show that speed and motor output are positively correlated with penetration and draft, and muckpile particle size is positively correlated with penetration and negatively correlated with draft (p < 0.0014).

By William H. Dresher

A recent report to President Reagan from the Business-Higher Education Forum, entitled, "America&apos;s Competitive Challenge: Need for a National Response," (1) observes that in the recent past, ". . . much of that research, which is necessary for major breakthroughs and sustained productivity improvements, was redirected to short-term objectives ... At the same time, American universities, which conduct more than half of the nation&apos;s basic research, have suffered from an accumulated under-investment in facilities and equipment amounting to an estimated $1-4 billion." Also contributing to the decline in America&apos;s productivity, the report said, are "...weak technology transfers between government and industry and between universities and industry..." This report as well as other recent reports and articles have, reopened a subject which has long been debated but rarely resolved-- the interaction between the academic community and industry in the sponsorship and conduct of research. The subject is unfortunately, too often treated as a controversial issue involving ethics and the profit motive when, in fact, the two branches of society -- industry and academia -- are necessarily part of the same system. I would like to speak to you today on the subject of university-industry interactions and to some extent discuss the role of government in such interactions. Finally, I will make some remarks as to how this subject applies to the mining industry.

Jan 1, 1984

By R. H. Grau

This Bureau of Mines report describes methane liberation trends from all U.S. coal mines liberating over 100 Mcfd (100,000 cfd) (2.8 Mm3/d). Detailed are results from the latest Bureau survey performed in 1985 and shown are mine and company name, location, coalbed and methane liberation quantity. The report shows that in 1985, there were 180 mines producing more than 100 Mcfd (2.8 Mm3/d) and the total quantity of methane liberated by these mines was 303,900 Mcfd (8,605 Mm3/d). This data is compared with previous studies performed in 1971 , 1973, 1975, and 1980. From 1971 to 1985, there was a 10% decrease in the number of mines liberating at least 100 Mcfd (2.8 Mm3/d) but there was a dramatic 34% increase of liberations from these mines. The increase seems to be caused by new mines being developed in deeper, gassier coalbeds, particularly in Alabama. Likewise, in comparing the ten highest liberating mines in 1975, 1980, and 1985, an increase in methane liberations each year was experienced. The report classifies mines by three levels of liberations: greater than 100 Mcfd but less than 500 Mcfd (2.8 to 14 Mm3/d); 500 Mcfd to 1 ,000 Mcfd (1 4 to 28 Mm3/d); and greater than 1,000 Mcfd (>28 Mm3/d). It was found that 87% of liberations were from mines producing over 1,000 Mcfd but these accounted for only 38% of all the mines in the study.

Jan 1, 1987

By Arash Nikvar-Hassani, Lianyang Zhang

To address the drawbacks of the conventional portland cement/fly ash (PC/FA)–based cementitious material currently used in practice to construct pumpable roof supports (PRSs), a novel cementitious material has been developed based on geopolymerization technology. The novel cementitious material is comprised of two pumpable grout streams: Stream 1 is a slurry composed of pozzolanic material and high-calcium additive and stream 2 is an alkaline solution comprised of a gelling agent. Biopolymer has been added to the second stream to provide ductility to the geopolymer cementitious material. In the laboratory scale, systematic experiments have been carried out to investigate the viscosity, setting time, and uniaxial compressive strength (UCS) of the new cementitious material and develop the right recipe for field applications. By collaborating with industry and the National Institute for Occupational Safety and Health (NIOSH), full-scale PRSs have been constructed and tested. The results show that (1) the novel cementitious material based PRSs has much higher peak UCS and similar or higher residual strength than the conventional PC/FA-based PRSs; and (2) unlike the PC/FAbased cementitious material, the novel cementitious material shows good long-term durability with nocracking/disintegration when exposed to air.

Jul 1, 2023

By J. F. Colinet, W. R. Reed, J. D. Potts

Dust surveys were conducted at six underground coal mines to determine if deep-cutting practices expose face workers to higher levels of respirable dust. The surveyed mines were able to successfully mine deep cuts without significantly increasing the dust exposures of face workers. In general, all of the selected mines exercised good dust control practices by maintaining water sprays, scrubber airflows, proper curtain setback distances and providing sufficient airflow to the active faces. For exhausting face ventilation, field data indicate that scrubber airflow is the most important factor for controlling dust. Data collected for this study indicate that 20-mesh screens should be cleaned for every 12.2 m (40 ft) of advance, because 22% of the deep cuts surveyed for this study experienced a 20% to 35% decrease in scrubber airflow over the course of the cut. For blowing face ventilation, field data and past research indicate that dust control is aided by maintaining a proper curtain-to-scrubber airflow ratio of 1.0 (measured before scrubber activation), as well as a curtain setback distance that allows the miner operator to stand at the mouth of the curtain for the entire cut. Dust levels on the bolting faces did not appear to be affected by the longer cycles associated with deep-cut mining when good ventilation practices were used.

Jan 1, 2011

By E. C. Watkins

The fiber glass industry is an industry that. hopefully is still in its infancy and far from a mature business. Fiber glass, as will be discussed in this paper, deals with continuous strand glass as used in the textile laminate and reinforced plastics industries. More specifically, we are speaking of glass used for printed circuit boards, tapes, paper, cloth, pipes, tanks, boats, recreational equipment, roofing shingles, and automotive applications. Compared to the flat, container and decorative glass industries, we are very young. The early patents on fiber glass applications were issued in the 1930&apos;s. The type of glass most commonly used in the world and that discussed in this paper is designated as "E" glass. This is a boro-silicate type glass having specific chemical and electrical resistance and low thermal expansion properties.

Jan 1, 1986

By J. E. Miller

The development of a new means of measuring the relative abrasivity of a slurry came about by necessity in 1967 during a full scale loop test of a 560 H.P. reciprocating pump handling magnetite slurry. In order to evaluate the life of expendable fluid end parts, the question of the amount of attrition in a re-circulated slurry had to be considered. An expected reduction in abrasivity as a result of this attrition would certainly give erroneous data on a parts life test. There- fore, some means of measuring such change in abrasivity became imperative. Several test methods were investigated1 but it was felt that these did not fully represent the wear due to back and forth rubbing action associated with a reciprocating pump. A "rubbing block" device used some years ago to determine the quality of hard facing materials was recalled, along with its shortcomings, in that an attempt was made to suspend a small percentage of abrasive particles in a colloidal viscous "mud". Erratic results were attributed to the fact that it was difficult to maintain a consistent slurry concentration under the metal wear block and the reciprocating speed was not great enough to maintain good agitation.

Jan 1, 1973

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

INTRODUCTION This manuscript is based on site inspections of five existing mines and the detailed analysis of two of the five. The detailed analysis and mathematical models of the two mines are presented herein in order to demonstrate a method of predicting mine water inflow using existing mines as a data base. Water management procedures were studied at all mines. Our original criteria for mine selection were that the mines be less than five years old and have an inflow rate in excess of 2000 gpm (7.6 m3/min). Background hydrologic and hydrogeologic data were required so that two of the mines could be used for our computer modeling experiments. Our original criteria proved to be difficult to meet. We contacted numerous mines via various sources. Many mines were unable to meet the cri¬teria stated or were not interested in cooperating with us. Therefore, we allowed some flexibility in the application of these criteria to satisfy our orig¬inal requirements. Although the visitations at all the mines are not discussed herein the characteristics of the two mines simulated are presented. These simulated mines consist of a uranium mine in New Mexico and a lead mine in Missouri. The mines and owners will remain anonymous by request. URANIUM MINE-CASE HISTORY NUMBER 1 Mine Location The uranium mine that we simulated is located in northwestern New Mexico. The ground surface at the uranium mine is at an elevation of approx¬imately 7000 ft (2134 m) above mean sea level. The climate is arid to semi-arid with an average annual precipitation of approximately 10 in. (25.4 cm) (John and West, 1963). Geologic Setting It is essential that a conceptual model of the stratigraphy, hydrostratigraphy, and structural geology of a mine site be developed prior to simulation by a mathematical model. If such a conceptual model is not developed prior to simulation major errors are difficult to avoid. The uranium mine we have simulated is located in the San Juan Basin (Fig. 68), which is a structural basin located in northwestern New Mexico and adjacent parts of Colorado, Arizona, and Utah. The sedimentary rocks present in the deepest part of the basin are nearly 15,000 ft (4572 m) thick (John and West, 1963). The regional dip of the sedimen¬tary rocks in the Grants area is generally northward toward the San Juan Basin; the regional attitude is modified locally by normal faults and folds (Hilbert, 1963). The rocks of interest in this area are sedimentary in origin and range in age from Penn¬sylvanian to Cretaceous. They rest on the Precam¬brian core of the Zuni Uplift. Associated intrusive and extrusive rocks are of Tertiary and Quaternary ages (Hilbert, 1963). The deepest sedimentary unit of interest at this mine site is the Morrison Formation as described by Hilbert (1963). The Morrison Formation con¬sists of three members in the general area of the mine site; the deepest member is the Recapture. The Recapture Member consists of alternating beds of gray sandstone and grayish red siltstone or mud¬stone with strata up to several feet in thickness. Only a few small deposits of uranium are contained within the Recapture Member. The Westwater Canyon Member overlies the Recapture Member of the Morrison Formation. The Westwater Can¬yon Member is approximately 150 ft (46 m) thick and consists of light brown to gray, poorly sorted, arkosic sandstone, and some interbedded gray mud¬stone. The Poison Canyon Formation denoted in subsequent discussions, tables, and figures was de¬termined to be a part of the Westwater Canyon Member. The separate connotation is used for the description of mining operations and hydraulic properties but a separate physical description is not included. This member of the Morrison Formation contains many large deposits of uranium. This is (some pages in this article are damaged)

Jan 1, 1986

By R. W. Smith

A number of investigators have studied the electrokinetic potential (?) of kaolinite as a function of pH. (1,2,3,4,5) These investigators all agree that the ? is negative except at quite acid pH values. The actual shapes of the curves reported by the several authors vary considerably. In spite of the general negative character of kaolinite the mineral has a rather high anion exchange capacity. (6,7,8) The cation exchange capacity is the lowest for the clay minerals. (6) It has been shown that anionic surfactants adsorb readily onto the mineral even up to neutrality, often in quantities exceeding the reported anion exchange capacity. (7,8) At least part of the anionic adsorption should be onto positive edge sites thought to exist on the mineral. (9,10) In the present study ? of kaolinite was measured as a function of pH using both the streaming potential (sp) and microelectrophoresis (m.e.p.) methods. In addition ? was measured at various pH values as functions of sodium dodecyl sulfate (SDS) and trimethyldodecylammonium chloride (TMDAC) concentrations. In some experiments decyl alcohol (DA) was added.

Jan 1, 1968

By R. H. Thorpe

In the earliest days of mining coal the obvious method was by bord and pillar working. Longwall in Britain developed from this as an expedient to overcome bottlenecks arising from ? 1) increasing depth of seams causing roof control difficulties; and 2) ventilation problems. This system became almost universal particularly in the thicker seams, where faces of up to one mile in length were not uncommon. Coal getting in these faces was usually by pick and shovel, filled into small waggons of up to 15 cwts capacity. These waggons ran on rails laid alongside the coal face and were often taken away from the face to the main haulage roads by horses. The roof was supported by rigid props and the wastes were normally controlled by strip packing. This system of mining was very flexible and worked well down to sections of about 3&apos;6"

Jan 1, 1976

By R. G. Reddy

Spent potliner is generated in the electrolytic process of reducing alumina to make aluminum metal using molten fluoride electrolytes. Spent potliner (SPL), obtained at the end of the pots life contains a small amount of cyanide and fluoride which present environmental problems. SPL became a listed hazardous waste under RCRA (K088) in July, 1990. One of the possible uses of spent potliner is as a substitute for Calcium Fluoride (CaF2) as a fluidizer in iron-and steel making and in the glass industry. A low viscosity and a low melting point are prime requirements for accomplishing the above. In the present study, viscosity measurements were performed. Effect of various additives, such as sand, limestone, dicalcium silicate dust, and MgO baghouse dust on the viscosity of SPL was studied. The results showed that with the addition of sand and limestone, very high fluidity of SPL melt can be obtained at significantly low temperatures. The results suggest that spent potliner mixed with additives could potentially be used as a substitute fluidizer.

Jan 1, 1999

By J. A. Dos Santos

"INTRODUCTION Dramatic reduction in energy use and environmental impact can be achieved with sandwich belt high angle conveyors that serve as the vital link in any IPCC (In-pit Crushing and Conveying) system. These high angle conveyors are not new at all, but have not found wide use in IPCC systems where they can realize the greatest advantage. Reduced energy consumption and environmental impact, with IPCC systems, was already realized in the 1970s. With the bulk material haulage limited to conventional open trough conveyors the most direct path out of the open pit was precluded, requiring low angle spiral ramps and/or deep slots and/or tunnels through the high wall of the open pit. These excavations, to accommodate the low angle limitations, represented undesirable impact on cost and on the environment. Against this backdrop, a major study in 1979 sought to develop high angle conveying systems that could continuously haul the mined bulk material directly out of the pit, along the high wall – the shortest distance between the two end points. Between 1979 and 1982, that study developed sandwich belt high angle conveying systems that utilized all conventional conveyor equipment including smooth surfaced rubber belts that could be continuously scraped clean. These systems had all of the positive features of conventional conveyors but overcame the angle limitation. By hugging the bulk material between two belts, the material’s internal friction could be developed to facilitate conveying at any high angle up to 90 degrees (vertical). After an intense testing period (about one year) on the first large scale prototype system, commercialization began in 1983 with the installation of a 60 degree incline system at a western United States coal mine, elevating 2000 t/h of coal to a train load out system. It did not take a long period of scrutiny and acceptance before this high angle conveyor found use in the most rugged requirements of an IPCC system. This was only the second commercial sale, and after more than 150 commercial installations, it remains arguably the most significant high angle conveying system. In 1984, a copper mine in Serbia, already using pit perimeter crushing and conveying, decided to move their primary crusher deep into the pit and to use a sandwich belt high angle conveying system to elevate the ore continuously, directly out of the pit, along the high wall to the pit perimeter where it then transferred to a conventional conveyor for the remaining haul to the plant. The system had significant features including 2000mm wide belts that elevated 250mm coarse ore, at 4000 t/h, over six 15-meter high benches for a total 90 meters of net lift. The system was able to reduce the truck haulage fleet by ten 200 ton trucks realizing great cost savings, zero emissions to the air and greatly reduced traffic congestion in the pit. The system operated successfully until 2002 when the mine shut down. Many successful sandwich belt high angle conveyors followed with the current count of commercial installations at more than 150. Despite the great success with this system, its use has not been repeated as part of an IPCC system."

Jan 1, 2018