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By E. R. Bauer

In response to the mandates in the MINER Act of 2006, the National Institute for Occupational Safety and Health (NIOSH) conducted refuge alternatives research that included characterizing the utility, practicality and survivability of refuge chambers and outby safe havens. NIOSH also prepared and delivered a report to Congress in late December 2007 that summarized the findings of the re-search, included recommendations concerning the design and performance specifications for refuge alternatives, and focused on specific information that could inform the regulatory process on refuge alternatives. This paper highlights NIOSH?s research and recommendations concerning refuge alternatives, survivability evaluations of refuge chambers and presents a brief review of the current deployment of refuge chambers in underground coal mines in the U.S. The research has lead to the conclusion that refuge alternatives have the potential for saving the lives of mine workers if they are part of a comprehensive escape and rescue plan and if appropriate training is provided.

Jan 1, 2009

By Taghi Sharizadeh, MAURICE SUNKPAL, ELAHEH TALEBI, PEDRAM ROGHANCHI

Excessive heat exposure is among the health and safety hazards that have adverse effects on the mine workers. The very nature of the underground climate makes it a convenient avenue for heat generation and accumulation from several heat sources including autocompression, equipment, and heat from strata. This study investigated the effects of clothing insulation and acclimation on the thermal comfort ofmineworkers who are exposed to heat in underground working environments. Several clothing ensembles have been considered to analyze the effect of the thermal resistance of clothing on the thermal comfort of workers. To investigate the impact of the clothing insulation, the thermal comfort limit of an underground miner was calculated in the form of maximum allowable exposure time for an 8-h shift. Parametric analyses were subsequently performed using an already developed thermal comfort model using MATLAB software. The results demonstrate that clothing insulation has a significant effect on the thermal comfort of mine workers. Increase in the clothing insulation decreases the maximum allowable exposure time of mine workers. This is especially true for the new and non-acclimated workers. Since acclimatized workers can tolerate a higher level of the dehydration and are capable of losing more water by sweat per shift, the maximum allowable exposure time for the acclimated workers are higher than non-acclimated workers. These analyses have shown that acclimated workers can have more continuous work per shift compared with non-acclimated workers. For example, in an environmental condition with clothing that a nonacclimated worker cannot work more than 5 h, an acclimated worker can work a full 8-h shift.

Jun 29, 2020

By Catherine C. Kilgore

The Bureau of Mines investigated the availability of fluorspar from 35 domestic and 50 foreign, mines and deposits in 13 countries. Only 6 domestic properties are considered to be economic reserves, as 29 domestic deposits were found to have an average total cost of production that was uneconomical at a breakeven (0 pct) DCFROR. The 50 foreign properties evaluated were selected to represent at least 85 pct of the known demonstrated fluorspar resources located in market economy countries and China. All evaluated foreign operations were found to be economic to marginally economic in January 1984 dollars. World demonstrated ore reserves of 344 million metric tons (mt) containing 123 million mt of CaF2 were evaluated for the 56 operations. Average total costs needed to cover all costs over the life of each operation were determined for all recoverable fluorspar products. Comparisons of mining, beneficiation, and transportation costs per mt of fluorspar concentrate illustrate advantages between countries. Of the three largest producing countries, China, with its high-grade, low-cost operations had the lowest weighted average total cost of production ($38/mt concentrate). Mexico was next at $54/mt concentrate, and South Africa needed an average of $90/mt of concentrate to cover all costs of production for its low grade deposits.

Jan 1, 1985

By R. O. Ifill

Alternating current electrocoagulation (AC/EC) may be an alternative to chemical coagulation in the thickening of tailings in the mineral, coal and petroleum industries. In the current study, ground silica (d50 as received, 20 µm) served as the model tailings solid with characteristically poor settling behavior. Its natural initial settling rate was 1.6 x 10-3 cm/min. This paper discusses the indirect application of AC/EC via aluminum electrodes to the enhancement of the settling behavior of the silica. ?Indirect AC/EC? means that the silica was coagulated upon dispersion in water that had been treated by AC/EC. The effect of AC/EC retention time on settling behavior is examined and the water chemistry is analyzed. Test results show that the initial settling rate of silica was enhanced by a factor of 4,000 at its maximum of 6.4 cm/min, under the conditions employed. Settling enhancement was achieved upon contact between the AC/EC-treated water and the silica, and the enhanced initial settling rate was dictated by the floc size. The aluminum dosage increased with AC/EC retention time, reaching a maximum concentration that was equivalent to 13.1% of that predicted by Faraday?s law. Complex aluminum ions are considered to have adsorbed onto the silica surface and an adsorption mass ratio of 0.3 mg Al/g silica was the most effective in enhancing settling behavior. Owing to pH change or slow kinetics, aluminum hydroxide precipitation did not lower the ion product ([Al3+][OH-]3) below the KSP of 6.3 x 10-32.

Jan 1, 2011

A group of geologists is located in a remote area of North America, without telephone lines, electricity or even running water. They have spent days accumulating data for a mining company that is eager to receive any information about the dig. But the team does not have access to a wired communications infrastructure. So the geologists must wait until they get to the nearest town to transmit their data. TMI Communications has a mobile satellite communication system that allows geologists in the field to transmit their data from remote locations. The company's satellite is the heart of its North American mobile communications network. This satellite turns all of North America, including Canada, Mexico and the Caribbean into a single digital communications cell. Traditional terrestrial communications systems, such as dedicated phone lines or cellular systems, cannot handle a remote geological exploration camp's extensive communications needs. Cellular technology, where available, has high costs and limited applications. For geologists in the isolated areas of North America, accessible and reliable phone service is vital to success. For those who have established temporary camps, TMI Communications' satellite phone service has improved the speed and accuracy of their communications with home base. This wireless network is suited to meet the needs of existing data applications and the challenges of new ones. It can support any one- or two-way, fixed or mobile data transfer. TMI's network enables geologists to transmit exploration results, spreadsheet information and location co- ordinates to anywhere in the world, at any time. This is important if a team needs to relay vital data back to its headquarters to support immediate decision making on a prospect site or a find.

Jan 1, 1999

By S. A. McGhee

Introduction ASARCO's Ray Operations include an open-pit copper mine in the Mineral Creek Mining District of Pinel County. The district is located in east central Arizona, 132 km (82 miles) southeast of Phoenix, AZ, and 126 km (78 miles) north of Tucson, AZ. The mine lies 8 km (5 miles) north of the Gila River in Mineral Creek Valley, between the Tortilla Mountains to the west and the Dripping Springs Range to the east. The rich copper deposits of the Mineral Creek mining district were first noted by a US Army Officer in 1847. The Ray Copper Company was the first major operation in the area. The company was named after the Ray mining claim, which was one of the earliest recorded claims. After decades of sporadic and unsuccessful mining ventures, the Ray Consolidated Mining Company was formed in 1910. The company introduced one of the first successful block caving operations in the world. In 1933, the Kennecott Copper Corporation acquired the Ray Consolidated Mining Company and continued operation of the mine under the same name until 1943, when the property was renamed the Ray Mines Division of Kennecott. Underground mining continued until 1955, when the mine was converted to an open pit operation. In November 1986, ASARCO Inc. purchased the Ray Mines Division and related facilities from Kennecott. The purchase also included the Hayden concentrator and the Ray Smelter, located approximately 32 km (20 miles) from the Ray Mine. This paper summarizes the 54,400 tpd (60,000 stpd) in-pit crusher and overland-conveying system and the 24,200 tpd (30.000 stpd) copper SAG concentrator and tailings disposal system constructed at ASARCO's Ray Complex - Ray Operations. General description of facilities The Ray concentrator includes a sulfide SAG concentrator, a "relocatable" in-pit primary-crushing plant and overland-conveyor system, a concentrator tailings-disposal system and plant ancillary facilities, The major elements of the project include: • A relocatable Fuller-Traylor in-pit 1524-mm (60-in) x 2261-mm (89-in) primary gyratory crusher at the 442 m (1450 ft) mine elevation. The crushing plant has a nominal capacity of 54,400 tpd (60,000 stpd) of sulfide ore. • A 1524-mm (60-in) wide x 1253-m (4110-ft) long overland belt conveyor, which runs from the in-pit crushing facility to a transfer station serving both the Ray Concentrator and the existing Hayden concentrator railroad load-out facility. • A 1524-mm (60-in) wide x 152-m (500-ft) long stackerbelt conveyor, which delivers a nominal 27,200 tpd (30,000 stpd) of ore from the transfer station to a 23,600 t (26,000 st) live-capacity coarse-ore stockpile serving the Ray Concentrator. • A nominal 27,200-tpd (30,000-stpd) concentrator, including semiautogenous grinding, flotation, copper concentrate filtration and associated plant ancillary facilities. • A tailings disposal system. Flotation tailings are pumped from the Ray Concentrator to a 122-m (400-ft) diam. thickener, located 4.8 km (3 miles) southeast of the Ray Concentrator site. The thickener overflow is pumped to two 76.2-m (250-ft.) diam. concrete process-water reservoirs. The thickener underflow, at 45% solids, is pumped to a new tailings disposal site located at Elder Gulch. • An HDPE-lined temporary-containment area to collect material from Ray Concentrator plant upsets and local storm-water run-off from a 100-year, 24-hour rain event. •A reclaim-water system from the Elder Gulch tailing pond area to the Ray Concentrator. • A 610-mm (24-in.) diam. x 31.9-km (19.2 mile) long fresh-water pipeline from the Hayden fresh-water-well field system to the Ray Concentrator. The overall process design criteria for the new Ray Concentrator is based on a nominal throughput of 27,200 tpd (30,000 stpd) of ore at 90% plant availability. However, based on the maximum performance capacity of the installed grinding equipment, it may be possible to achieve a sustained operating level in excess of 36,300 tpd (40,000 stpd) when processing soft ores. The balance of this paper will cover the major process systems including:

Jan 1, 1995

By A. Anani

The goal of hot mix asphalt (HMA) design is optimal material blending to meet Superpave specifications. HMA cost minimization has being modeled as linear programing (LP) (Awuah-Offei et al 2011) and mixed integer LP problems (Brown et al, 2012) and solved with commercial software such as LINDO and CPLEX. Often these software, which are designed to solve general optimization problems, have licenses that cost thousands of dollars. The objective of this study is to develop a novel solution algorithm to the HMA optimization problem presented by Brown et al. (2012), which will negate the need for expensive commercial packages and is specific to this problem. This novel algorithm is based on the branch and bound algorithm for solving mixed integer problems. The algorithm is incorporated into a software package with an easy-to-use graphical user interface. The developed algorithm is validated using a real-life HMA problem. Using this solution algorithm will save companies thousands of dollars, otherwise used to purchase software licenses.

Feb 23, 2014

By J. Mark Richardson

The use of an integrated software package to simulate and optimize the performance of mineral processing circuits is illustrated by three examples. Each represents application of the software and associated techniques to optimization of an existing plant. In the first example, a South-western U. S. copper company employs the method to simulate and improve performance in crushing, screening and SAG milling. A second copper producer evaluates equipment changes and flowsheet alternatives in a closed-circuit ball mill-cyclone operation. A Southeastern U. S. aluminum refinery optimizes performance of a dry milling and classification circuit for production of carbon particles used in anode casting. Graphical and tabular data describe the circuits simulated and the results achieved in each example.

Jan 1, 1990

By John H. Bassarear

Large scale equipment can reduce both capital and operating costs for concentrators. Large semiautogenous mills can be effectively utilized in big mills but they must be kept fully loaded to take advantage of their great potential. With varying hardness ores, controlling throughput is difficult and careful design must take this into consideration. A variable speed drive on large primary semi- autogenous mills can overcome this problem to a considerable degree. This paper provides a base case using standard accepted technology of two wound rotor single speed 4,476 kW (6,000 hp) motors to drive the primary SAG mills of a hypothetical 36,364 mtpd (40,000 stpd) typical porphyry copper mill. One other fixed speed drive will be described and five variable speed drive configurations will be reviewed. Capital costs for the entire grinding circuit, from crusher product to flotation feed, were developed for the base case and each of the alternatives. A technical evaluation comparing the advantages and disadvantages of each drive is presented.

Jan 1, 1985

By Madan M. Singh

A new mining project can raise questions about what effect it will have on the community after it closes. Historically, little was done by communities to plan for and accommodate the changes in the economic infrastructure when a mine project is finished either after the ore runs out or if market conditions deteriorate and make it unprofitable. Projects that were underfunded were generally unable to withstand periodic market downturns. However, data in a recent study conducted by the Seidman Research Institute leads to the conclusion that, even when a mining project is terminated, there can be (and, in the specific case studied, is) a residual beneficial economic effect in the community, which lasts for a long period of time.

Jan 1, 2010

By John Hillman

The reconstruction of the British mining industry which began shortly after nationalisation in 1947 continued until well into the 1950s with the support of the Government. This support diminished as plentiful supplies of cheap oil from the middle East made many coalmines un¬economic. The planned rundown which followed was foreshadowed in the 1965 White Paper which outlined the financial measures to speed the closure of uneconomic collieries. The decline appeared likely to continue into the 21st century as the role of coal in UK energy requirements became less and less significant. This policy was well on course in 1973 when the oil crisis led to a complete reappraisal of the position. A Government examination in 1974 of the coal industry then supported the National Coal Board's general strategy of major capital investment as set out in its 'Plan for Coal' (1). Plan for Coal aimed at halting the intended decline in the industry's capacity and restoring it to 120 million tons of deep mined output in 1985. The Plan proposed action along the following lines (a) increase output by major Investment in construction at existing collieries with sufficient reserves to give 13 million tons per annum extra output (b) invest in extending the life of existing collieries by gaining access to additional reserves, to give 12 million tons per annum extra output © by sinking new mines, produce an extra 20 million tons per annum. Figure 1 shows the contribution (2) expected from each of the three sources up to 1990/1. The measures taken were expected to offset the inevitable exhaustion of existing pits and provide the additional 40-42 million tons capacity needed to restore annual output in 1985 to 120 million tons. Parallel to this programme for expansion of deep mined output, opencast mining was to produce 10-15 million tons per annum. A recent review however, indicates that new mines are likely to yield only 10 million tons per annum by 1985, mainly owing to the need for public inquiries into proposals for new mining and the resultant fact that planning approval takes 'much longer than originally expected. Nevertheless the objective of 120 million tons deep mined output by 1985 is still considered achievable of means of a greater contribution from existing mines. This programme was originally estimated to cost 2,-56 million dollars at March 1974 prices, out revised estimates of the cost of Plan for Coal indicate that it may now exceed 6,960 million dollars at mid-1977 prices. Plan for Coal includes a substantial investment in new and reconstructed coal preparation plaints. At mid-1977 prices more than 322 million dollars Ls so be invested in coal preparation projects in the UK, at the rate of, about 70 million dollars a year. This rate 3f capital expenditure has been maintained over the past three years and covers investment in a wide range of coal preparation coals for the power station, coking and domestic markets.

Jan 1, 1978

By P. C. Dagneau

QUEBEC CARTIER MINING COMPANY was incorporated in 1957 to explore and develop low grade iron are deposits in Northeastern Quebec. Four separate deposits were located and identified as being of ore quality and size to support commercial development. Said deposits were located at Lac Jeannine, Mount-Wright, Fire Lake and Mount-Reed. In 1970, in preparation for the replacement of the depleting Lac Jeannine are reserves, construction of an SS-mile railway extension was initiated, towards the newly selected mine site at Mount-Wright. This orebody was selected because the physical and chemical characteristics of the ore were very much similar to those of Lac Jeannine, as well as its huge reserves, estimated at more than one billion tons. In addition to the railroad extension, a mine, concentrator and related facilities were constructed and the new town of Fermont was created. Total investment for the Mount-Wright project exceeded $700 million and initial production commenced in 1976. Quebec Cartier is presently (1992) the leading producer and seller of iron ore in North America at 15,5 million tons and employs some 2,300 people.

Jan 1, 1993

By T. W. Martin

In 1973 the percentage of total coal production mined by surface methods in the United States again increased, reflecting a long term trend. Barely three decades ago less than 10 percent of the U.S. coal was produced in this manner. Last year surface coal mines produced 299 million tons, 50.7 percent of our nation's coal. At present there are about 2450 strip mines producing coal in 23 states. While accident rates for surface coal mines are less than underground coal mining rates, surface coal mining is certainly one of the more hazardous occupations. Figure 1 shows the injury rates (fatal and non-fatal accidents) of major U.S. industries. Note that surface coal mining is more hazardous than all other industries except contract construction. Accident rates are about the same as the contract construction industry which uses similar equipment. A limited amount of detailed data was available on elements of surface coal mining jobs and this lack of compiled information made it difficult to assess the significance of the conditions contributing to accidents and to make recommendations that would reduce accident frequency. Therefore as part of the Bureau of Mines continuing research efforts related to the health and safety of the coal miner, a contract was awarded to Theodore Barry and Associates of Los Angeles to conduct an "Industrial Engineering Study of Hazards Associated with Surface Coal Mines."

Jan 1, 1974

By Chen-Luh Lin, Jan D. Miller, Sindhoora Puvvada

Most of the future phosphate resources in Florida, which accounts for 70  percent of production in the United States, are contaminated with too much dolomite that is difficult to remove with existing industrial practices. The low-grade ore quality coupled with increased phosphate demand is of concern to the Florida phosphate industry. In order to resolve the dolomite issue, various beneficiation technologies were designed and evaluated, but none have been sufficiently successful [1]. 

By J. R. Goode

The first part of this paper comprises a general discussion on refractory gold ores and their mineralogy processing methods now in use are reviewed and test procedures for the important methods are discussed. Factors determining the selection of a process are outlined. The paper includes a survey of production plants using pressure oxidation, roasting and biological oxidation to process refractory ore. The status of processes under development is briefly reviewed.

Jan 1, 1993

By Weijie Wei, Jinwang Zhang, Shengli Yang

Due to the unbalanced top coal drawing process, the recovery ratio is low, and the equipment stability is poor in the longwall top coal caving mining (LTCC) with an inclined seam. Aimed at these problems, this paper proposes the dynamic group caving method (DGCM). The basic procedure is that when the drawn volume of top coal from a support reaches about half, the next support begins drawing top coal, and the drawing process between the two supports is half done. Taking the panel no. 8103 in Beixinyao mine as the research object, the characteristics of drawing body shape, particle movement trace, and variation of recovery ratio are investigated by using Particle Flow Code. Furthermore, the effect of mining-caving ratio on the top coal drawing law under DGCM is revealed by the simulation results, and compared with the single-opening sequence caving method (SSCM) and the doubleopenings sequence caving method, advantages of DGCM are determined in inclined seams. The results show that the drawing body can be divided into the left and the right under DGCM. Specifically, the drawn top coal particles in the left are from the upper part of this support, that in the right are from the lower part of next support, which verifies the correctness of the theoretical analysis. The characteristics of the drawing body shape are more significant when the mining-caving ratio is relatively small. In addition, with increasing mining-caving ratio, the recovery ratio of the panel both increases gradually underDGCMand SSCM.When the miningcaving ratio is less than 1:1, the recovery ratio of DGCM is obviously higher than that of SSCM, indicating that DGCM can greatly improve the recovery ratio. While when the mining-caving ratio is larger than 1:1, the recovery ratio of two caving methods is basically equal; however, the advancing time is significantly shortened under DGCM. This study are of great significance to realize high efficiency and precise caving in LTCC with an inclined seam.

Jul 7, 2020

By Muhammad Azeem Raza, Muhammad Zaka Emad, Muhammad Usman Tahir, Saqib Ahmad Saki, Muhammad Usman Khan

Shotcrete is a key initial rock support component used in underground mining and civil projects. The strength of shotcrete plays a vital role in providing and deciding the support adequacy. The strength of shotcrete is evaluated by its uniaxial compressive strength (UCS) test that is usually conducted on extracted core samples. Shotcrete cores are commonly extracted from the test panels along the direction of its application; thus, its strength perpendicular to the direction of application is often neglected. The true bearing strength of shotcrete is thereby ignored despite the fact that higher stresses act normal to the direction of its application. Although the anisotropic behavior of shotcrete is well reported, no study has reported extensive testing of this scale and nature. This study reports extensive testing of shotcrete cores extracted from tunnels of a mega hydroelectric project in Pakistan. The study reports 7- and 28-days strength tests of shotcrete cores. The research also investigates the anisotropic behavior of steel fiber reinforced shotcrete and suggests that shotcrete exhibits higher strength perpendicular to the direction of its application.

Dec 26, 2022

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 Ken Deter

I have some general comments on optimizing profitability based on working in heap leach operations in tropical environments. Many of my planned comments this morning have been stated earlier in the Conference. I get the opportunity to say only a few words, and follow up on some specific comments and concerns. In normal operations, stack the ore higher, where possible. The surface area of the stacked ore controls the total solution volume that can be sprayed on the pad at a specific time period. With higher lifts, solution will contact more ore particles as it percolate through the lift, and therefore the solution grade flowing through the base will be increased, whether it is freshly stacked or almost completely leached ore. This also allows more tonnage to be stacked per lift on the pad, decreasing conveyor movement and downtime. Increasing from 6 to 8 meter (20 to 26 feet) lifts have worked successfully, and some mines are using 10 meter (33 feet) lifts in 10,000-15,000 dmtpd plants. The same concept will work with smaller plants, but initial 6 meter lifts should be used until the start-up period has shown the leaching is working as planned. As the stacked ore lifts get higher, the solution flowing through the highest lift is diluted by solution retained in the lower ones. Eventually no real increase is noted from putting fresh ore in service, and gold production rate suffers. To address the dilution problem, an HDPE liner should be installed between each 2 to 4 lifts, depending on each site specific lift height, recovery, % slump in the lift, recovery rates, and grade treated, among others. The surface of the previous lift should be leveled and sloped to control the effluent draining from the planned lift, then the liner is placed, including berms and discharge pipes. The intermediate HDPE liner does not need to be completely sealed as would be the case with the base liner, and can accept some solution penetrating into the lower lifts. A one-mm (40-mil) thickness liner can be used. This intermediate liner allows the upper lift(s) higher-grade effluent to report to the pad discharge point. This will maintain higher gold solution grades going to the ponds. The slopes of the previous lifts should not be covered with HDPE in this concept, and precipitation falling on the lower lift slopes and benches can be directed to barren or surge ponds.

Jan 1, 1999

By Smart Chika Obiora, Anthony Chukwu

Mining and processing of raw materials is one of the earliest activities of man for survival and development. The products are used for making jewelleries, metal tools, and for exchange of goods and services. However, despite the benefits of mine recoveries, it’s negative and long term effects on the environment and lives remain the point of discussion throughout the world. This study summarizes findings of researches in active/abandoned lead–zinc mines in Enyigba and Ishiagu mining districts including social surveys in the localities surrounding the mines in order to ascertain the effects on the inhabitants. Lead–zinc mining in Enyigba and Ishiagu, southeastern Nigeria has lasted for decades and has affected the socio-economic well-being of the host communities. There are evidences of soil, water, and plants contamination by Pb, Zn, Cd, and Mn in these mining communities. The mining activities also affect the overall social development of the host communities. These effects include retardation in educational development, health care provision, unemployment, and access to potable water supply. Others are rise in poverty level, social amenities, sanitary/hygiene level, high cost of living, and fertility rate. Possible mitigation strategies such as to liaising with relevant government agencies in collaboration with mine owner, mine workers, and host communities can ameliorate these negative effects and promote environmental sustainability.

Feb 24, 2023