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By J. Steenkamp, A. L. Hannaford, A. N. Jones

O’Okiep Copper Co. Ltd.’s 45-kt/month (49,600-stpm) concentrator at the Spektakel mine in South Africa is comprised largely of reconditioned equipment from the old O’Okiep mill, which operated from 1940-1975. When combined with modern approaches to layout, crushing, classification, and flotation, a simple-to-operate plant was constructed at a cost of only $3.2 million.

Jan 11, 1982

By Robert L. Frantz

The U.S. Bureau of Mines (USBM) established on August 15, 1983, a Generic Mineral f Technology Center (GTC) for Respirable Dust within the Mining and Mineral Resources Research Institutes (MRIs) at The Pennsylvania State University (PSU) and West Virginia University (WW) in association with participating MRIs at Massachusetts Institute of Technology (MIT) and the University of Minnesota (UMN). The primary goal of the Center is to reduce the incidence and severity of respirable dust disease through researching and advancing the fundamental under- standing of all aspects of respirable dust associated with mining and milling and the interaction of dust and lungs. Other Center activities include the training of engineers and scientists, graduate students and undergraduate students through their respective institutions, and the technology transfer to the industry. The Center also serves as the reference center for publications in the respirable dust area.

Jan 1, 1986

By R. C. Moe

Pelletizing studies were conducted to find ways of improving the metallurgical properties of acid and fluxed pellets. In acid pellets, melting temperature was controlled by the formation of low melting temperature fayalite during reduction. Formation of a permeable metallic shell during reduction allows diffusion of reducing gas to the pellet core, decreasing fayalite formation. With fluxed pellets, slag chemistry had the main influence on metallurgical properties. Mineralogical, chemical and structural transformations that occur during reduction were studied at various intervals for a number of Mesabi Range pellet types. Six pellet types were selected for further investigation based on significant differences in softening/melting characteristics. Retention of porosity in the metallic shell was found to be the overriding factor in determining softening/melting behavior.

Jan 1, 1990

By C. J. Hunter

Prior to the commissioning of the Sansu Sulphide Treatment Plant (STP), Ashanti Goldfields' treatment facilities were capable of producing some 700 000 ounces of gold per annum. Ore sources were underground refractory sulphides, surface oxides and old tailings dams. In order to increase the mines production to 1 million ounces per annum and to accommodate the change in ore sources as surface mining moved into fresh sulphides, a part of the strategy adopted was -the construction of a new sulphide treatment plant with a capacity to mill 2.2 million tonnes per annum and produce 400 000 ounces of gold. The STP has been designed, built and commissioned. and has been in operation since February 1994. The Plant boasts the world's largest Bio-oxidation facility which is capable of treating 790 tonnes of sulphide concentrate per day at a grade of 11.4% sulphur and is currently undergoing an expansion to increase its capacity to 1050 tonnes per day. The following sections describe the reasons for selecting the BioX® Process, the plant design, the commissioning experience and the plant performance to date.

Jan 1, 1996

By Bastiawarman Rifki, Nugraha Nico, Priatna Anton

The Deep Mill Level Zone (DMLZ) is one of the underground mines operated by PT Freeport Indonesia (PTFI). DMLZ is located approximately 1,500 m below ground level, where dominant rock types are skarn and diorite. The DMLZ mine and layout was designed based on the previous mine located above ground, the Deep Ore Zone (DOZ). The DOZ mine, located approximately 1,000 m below ground level, was a proven success, with a maximum capacity of 80K tonnes per day. During the production ramp up, which started in 2015, the DMLZ experienced several seismic events. The most significant event occurred in January 2018 with a recorded magnitude of more than 2.0 Mw, which resulted in a 14-month delay of the undercutting until March 2019. The significant event in January 2018 required the DMLZ to apply several adjustments related to the mining plan to help ensure it could continue development and achieve production target safely. This paper will describe the changes in the DMLZ mine plan as a response to the seismic activity at the mine, the reason behind the decisions being made to change the mine plan and strategy, as well as strengths and weaknesses between the previous plan compared with the new plan.

Apr 26, 2022

By James A. Brierley

The author has known Professor Wadsworth since the 1960s, when interest in the application of microbial technology for metallurgical processes was beginning its rapid rise. The development of bioleaching and mineral biooxidation was advanced, in part, by Wadsworth's belief in the potential for its application, as well as by his numerous contributions to the understanding of the fundamentals of hydro-metallurgy. Wadsworth is recognized for fostering research and development of the field and adding credibility to microbiological-based metallurgical processes, now often referred to as biohydrometallurgy. This Wadsworth Award is recognition of the field of biohy-drometallurgy. It is an honor to serve as a representative for all those individuals involved in research and commercial development of this technology. Modern commercial application of biohydromctallurgy began with the bioleaching of copper from submarginal-grade. run-of-mine material. Kennecott Copper successfully developed this process in the 1950s at the Bingham Canyon Mine in Utah. Other mining operations around the world followed Kennecott's lead. Today, dump bioleaching remains a very low-cost process for scavenging copper from rock that cannot be economically processed by any other method. Despite the commercial success of dump bioleaching, little effort has been made to enhance the micro Biological component of this leach process. Expansion of biohydro metallurgy into recovery of other metals did not occur until the mid-1980s, when the first commercial plant for the pretreatment of refractory gold-bearing concentrate was commissioned at the Fairview operation in south Africa. Now, processes for copper bioleaching and refractory gold pretreatment are engineered specifically to promote the activity of the microorganisms. This paper reviews the current state of commercial applications of biohydrometallurgy. It examines potential future developments and portrays the role of the microbiologist in facilitating continued growth and commercial acceptance of the technology. The commercia1 examples in this paper illustrate the expansion of microbial processing in the mining industry during the last 20 years. Copper bioleaching Early commercial applications of bioleach technology involved processing submarginal-grade, copper- bearing, run-of- mine rock in dumps. Recent applications of the technology use engineered bioleach heaps. Currently. all of these operations are in the Southern Hemi-

Jan 1, 2000

By Vincent Tricot, Daniel Altier, Roger Storry

"INTRODUCTION The Liantang / Heung Yuen Wai Boundary Control Point (BCP) project, located in the north-east New Territories of Hong Kong, will form the seventh land crossing between the Hong Kong Special Administrative Region (HK) and Shenzhen in China (Figure 1). The BCP project comprises an 11km long dual two-lane trunk road linking the border crossing facility to the existing road infrastructure in HK and when completed it will facilitate future regional development and alleviate frequent traffic congestion at the existing Boarder control points.In December 2013 Dragages Hong Kong (a member of the Bouygues Construction Group) was awarded the 5.4km long Contract 2 section of the project, including the 4.8km long Lung Shan Tunnels, by the Civil Engineering and Development Department of the Government of the HK SAR. To overcome the difficult ground conditions anticipated along the Contract 2 section of the project the tunnels will be constructed using mechanical and drill and blast (D&B) excavation techniques together with a 14.1m diameter multi-mode Earth Pressure Balance Tunnel Boring Machine (TBM). Land constraints imposed on the project alignment require that the final 500m section at the northern end of the Lung Shan tunnels, under Princess Hill (Figure 2), takes an almost 90 degree turn towards the north. The resulting 354m radius curve restricts the sight-line of road users to below the minimum 110m at 80km/hr hence, the internal span of the tunnels needs to be increased from 12.6m, in the standard straight section, to 15.4m in this curved section to accommodate a widened shoulder which will provide the required sight-line distance."

Jan 1, 2016

By W. W. Wen

Approximately 3 percent of the raw coal processed in U. S. Steel coal-preparation plants ultimately reports to slurry- tailing disposal. The usual disposal method is to pump the refuse slurry containing the minus-48 mesh (300 µm) refuse into dams or impoundments that are constructed of the coarse refuse. The failure of an impoundment at Buffalo Creek, Saunders, West Virginia and the sliding failure of a coarse refuse pile at Aberfan, Wales, England, resulted in an evaluation of refuse disposal methods by various regulatory agencies as well as coal producers. The result of this evaluation was the promulgation of Federal regulations dealing with the construction, operation, and final abandonment of coal refuse disposal areas. As a consequence of these regulations the use of the landfill/impoundment method of disposal became in some cases economically urlfeasible or technically impossible. U, S. Steel has studied a number of alternative methods of refuse disposal. During an investigation of mechanical dewatering with a pressure filter, Portland cement was added to the slurry to solidify the filter cake. Because of the success of the cement addition it was decided to study the stabilization of fine-coal-refuse slurry through use of cement-type additives.

Jan 1, 1979

By K. J. Shoop

An analysis of an iron ore processing plant showed that its performance was affected by seasonal temperature variations, resulting in reduced grinding efficiency in the winter months. Laboratory and pilot plant studies were conducted to determine the cause of the observed loss of efficiency at lower temperatures. Experiments were divided into two groups: 1) Drop testing, which was performed to determine the effect of brittleness and frozen pore water on the breakage characteristics of the ore. 2) Wet grinding, which was further divided into laboratory tumble tests and pilot plant scale tests. Laboratory tumble tests were performed to determine the effect of temperature on the grinding of iron ore. This was done by varying the percentage of fines and percentage of metamorphic clay material added to the iron ore grinding charge. The pilot plant scale tests were conducted as a scale-up of the laboratory tumble tests and to determine the effect of temperature on slurry rheology, and how this changing rheology affected the grinding of iron ore. The combined results of drop testing and wet grinding were used to make some predictions of the performance of plant scale autogenous grinding. In drop tests the breakage characteristics of the iron ore had no noticeable change in size stabilities for dry and water saturated samples over the temperature range studied (-25°C to 25°C). The effect of frozen pore water was not observed due to the low apparent porosity of the iron ore. The addition of fines to the test charge caused a decrease in the amount of new fines produced for both the laboratory tumble tests and pilot plant scale tests. The production of fines further decreased when tests were performed at the colder temperature. The increased energy consumption at lower temperatures was dominated by rheology phenomena. The addition of metamorphic clay material to the iron ore test charge resulted in more fines being produced in the laboratory tumble tests. This was due to the metamorphic clay material grinding easier than the iron ore. This caused an increase in the amount of fines in the mill, which in turn, would prematurely slow down the grinding of iron ore because of an increase in the viscosity of the slurry.

Jan 1, 1995

By K. J. Reid, I. Iwasaki, T. Sabelin

Introduction Occurrences of platinum group minerals (PGM) in mafic and ultramafic rocks in association with Cu-Ni mineralization are well known. Yet for the Duluth Complex, a large body of dominantly mafic igneous rocks in northeastern Minnesota with significant Cu-Ni mineralization (Bonnichsen, 1972; Boucher, 1975; Watowich, 1978; Weiblen and Morey, 1980), little has been published concerning its platinum group element (PGE) contents and mineralogy. In a previous study of massive sulfide rock, sperrylite (PtAs2) was found in nickel arsenide and recovery of arsenide minerals by flotation has been investigated (Iwasaki et al., 1986; Ryan and Weiblen, 1984). Recently, an announcement was made of the discovery of significant quantities of platinum and palladium [peak assay values of 9 g/t (0.27 oz per st)] in an interval of a drill core with disseminated sulfide mineralization 16 km (10 miles) southeast of Ely, MN (Anon, 1985; Sabelin, 1985). This paper describes PGM from the Pt-Pd-enriched zone and evaluates their potential for concentration. Geologic setting The Pt-Pd zone occurs in the South Kawishiwi in¬trusion near its basal contact at 734 to 737 m (2408 to 2418 ft) depth. The South Kawishiwi is one of several minor mafic intrusions that constitute the western margin of the Duluth Complex near Ely, MN (Fig. 1). It is com¬prised of a 70.1- to 350.5-m (230- to 1150-ft) thick sul¬fide-bearing basal unit of troctolite with minor amounts of anorthosite, norite, picrite, and oxide layers overlain by a troctolite unit with interlayered anorthosite (Bonnichsen et al., 1980; Foose, 1984; and Foose and Weiblen, 1986). The basal unit, in which the Pt-Pd zone is found, is markedly heterogeneous in rock type, lacking any coherent internal stratigraphy, and shows evidence of both pre- and post-consolidation faulting (Foose and Weiblen, 1986). Geochemistry Chemical analyses were made on 0.3-m (1-ft) intervals of crushed drill core material (1/4 core, BX size) at 731.5 to 740.7 m (2400 to 2430 ft) depth and the results are presented in Fig. 2. The analytical techniques used were atomic absorption spectrometry for silver, chromium, copper, and nickel; a coulometric titration method for sulfur; and a combined fire assay-DC plasma technique for gold, palladium, and platinum. Figure 2 shows the significant enrichment of platinum and palladium at 734 to 737 m (2408 to 2418 ft) depth; it corresponds to a stratigraphic interval rich in oxide minerals. Average values for platinum and palladium contents over the 3-m (10-ft) interval are 2520 and 2580 ppb, respectively, or 5.1 g/t (0.15 oz per st) platinum plus palladium. Platinum plus palladium contents from individual 0.3-m (1-ft) intervals are as high as 9120 ppb or 9.1 g/t (0.27 oz per st) and average 7.4 g/t (0.22 oz per st) in a 1.5-m (5-ft) interval in the richest part of the zone. The Pt-to-Pd ratio is nearly 1:1 over the entire 9.1-m (30-ft) interval and the platinum and palladium correlate closely with chromium, which ranges from 1.9% to 5.3%. Copper, nickel, and sulfur contents for the drill core interval are also given in Fig. 2. In¬terestingly, the Pt-Pd high corresponds to a depletion in copper, nickel, and sulfur. Mineralogy of the Pt-Pd zone The Pt-Pd zone occurs within an oxide-rich zone or layer that consists essentially of oxide minerals, plagioclase, and olivine. The oxides may constitute as

Jan 1, 1987

By J. N. Economopoulos

The mining industry is becoming increasingly aware of environmental issues and has made big forward steps in the field of environmental management. Legislation, social pressure and competition introduced changes in the industry?s policies and practices in order to conform to the new standards of environmental performance. The paper addresses the need for an integrated and systematic approach to the issues of environmental management within the mining industry.

Jan 1, 2002

By Larry S. Diaz

It is felt that prior to the presentation of the technical portion of this paper, some general information would be of interest. The Sar Cheshmeh Copper Mining Company was founded by the Iranian Government to oversee the design and construction of Iran's first copper smelter complex, which will be another step in the industrialization of Iran. Since this is the first such plant in Iran, Sar Cheshmeh wisely chose to enlist the technical assistance needed in regard to the facilities requiring engineering and operating know-how from outside of Iran. To obtain the foregoing, SCCM signed a contract with Anaconda-Iran, Inc., which was formed by The Anaconda Company for this project. The general scope of this contract covers the overseeing of engineering and design by Parsons-Jurden as well as physical plant construction. In addition to this, the same services were given to water supply and tailings disposal systems which were not included as part of the Parsons-Jurden contract. Anaconda-Iran is also assisting SCCM in setting up a training program for Iranian personnel for plant operations, administration, etc., which .will go on for some time after the plant startup, under the guidance of Anaconda-Iran expatriates. SCCM?s other responsibilities, in respect to the development of this industrial effort, are the building of a townsite for approximately 2200 families, the expected manning requirements, having all the facilities necessary to provide a pleasant life in a remote area. It has been estimated that the manpower for operations will be 2050.

Jan 1, 1976

By R. L. Grayson, R. C. Althouse, L. A. Layne, M. J. Klishis

Introduction The researchers combine a detailed microanalysis of accident data and roof bolting field observations to develop new and meaningful risk indices for the roof bolter job classification. These indices incorporate information that has not been used in traditional measures of risk. The significant parameters that these new indices incorporate are the total number of accidents occurring within a particular work activity, the amount of time a miner usually spends performing the work activities in which the accidents occurred, and the average severity of the injuries incurred in the accidents. Being specific for a particular job, they aid comparisons among work tasks performed in the bolting cycle. The indices' specificity provide anew basis for reexamination of training and operational development of miners for specific jobs. Microanalysis The roof bolting microanalysis was performed by examining the empirical data and written summaries for all bolting accidents reported in the West Virginia Safety Information System (WVSIS) for the years 1983-1987. The WVSIS database contained more than 15,000 accident narratives. The microanalysis of bolting accidents covered 2083 reported personal injuries (RPIs). This large number of injuries, performed at an industry-wide level, provides a broad picture of the nature of roof bolting injuries. The analysis focuses on the activities associated with the cyclical routines of the roof bolting process. For this analysis, roof bolting tasks were categorized into work routines based on the sequence of work activities performed by roof bolters. Four work routines were identified within the overall bolting cycle. And these routines comprised 70% (1217) of all roof bolting accidents at the face. The routines include: face area preparation; tramming, positioning and setting the automated temporary roof support (ATRS); drilling holes; and installing bolts. The types of injuries most common within a routine and the average severity of those RPIs were also identified. Table 1 gives a breakdown of the percentage of accidents by the bolting routines. Another 19% of the roof bolting RPIs cannot be classified in a particular routine because of the vagueness of the narratives. However, it was clear that these accidents involved drilling holes or installing bolts. These results are consistent with previous research. It identified drilling and bolting, followed by tramming, as the most frequent sources of injuries (Helander, Krohn & Curtain, 1983). [Table 1-Accidents by bolting routines at the face Accidents Bolting routine Percent Number Face area preparation 18.2 222 Tram, position, ATRS 23.3 284 Drilling holes 34.2 416 Installing bolts 24.2 295 Total 1217] Roof bolter operators were involved in 82% (1706) of the 2083 accidents. The RPIs incurred by classified roof bolters occurred largely on working sections (95%) and at the face area on the working section (75%). Approximately 23% of the 1706 roof bolter RPIs occurred during tasks which were not part of the bolting cycle. These included belt utility work, providing supplies and section moves. Miners in other job classifications who were involved in roof bolting accidents accounted for the remaining 18% of the 2083 roof bolting RPIs. Among this group were section utility men and general laborers (41%), section machine operators (38%), and electricians and mechanics (11%). Another 9% were performing on-section supervisory functions. Face area preparation Face area preparation work led to a substantial number of injuries, although it accounted for the lowest number of injuries of the four routines. Several tasks were associated with the injuries incurred in this routine. Table 2 shows the percentage of injuries associated with face area preparation. [Table 2- Accidents occurring during face area preparation 26.6% Setting temporary and supplemental supports 24.8% Scaling roof and ribs 21.2% Installing ventilation devices 15.8% Handling supplies and materials 11.6% Miscellaneous tasks] The injuries commonly involved overexertion and 'struck by' falling coal and rock. Handling materials often led to overexertion injuries. Nearly 66% of the scaling-related injuries and about 33% of the injuries incurred during the setting of temporary supports involved 'strikes' by equipment, materials, top and rib. Tramming, positioning and setting the ATRS During tramming operations, workers were often struck by components of the machine and hit by loose roof and rib. Poor footing also led to slips, sometimes against the moving machinery. Table 3 gives a breakdown of accidents in this routine. Control levers were a significant problem in this routine.

Jan 1, 1993

By L. C. Lang

INTRODUCTION The introduction of 165-mm (61/z-in.) holes to underground mining operations has made possible the application of Canadian Industries Ltd's (CIL) vertical crater retreat (VCR) mining method, illustrated in the accompanying sketches. This unique and revolutionary new application of spherical charge technology (see the Appendix), when applied to primary stopes and pillar recovery, eliminates raise boring, slot cutting, and dilu¬tion of ore by backfill; greatly improves fragmentation; reduces labor and time requirements; eliminates uphole drilling and blasting; and minimizes or completely elimi¬nates damages by blasts to the walls and retreating backs of the stope or pillar. If vertical large diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at the calculated optimum distance from the back of the undercut and detonated, a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, min¬ing of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed. The VCR method is also applicable to drop raises and has the potential to replace all other raising methods under most circumstances. PILLAR RECOVERY Levack Mine Inco Metals Co., Ontario Div., provided the first opportunity for the method in pillar recovery. The Levack mine's high grade ore pillar No. 4800 on the 975-m (3200-ft) level was used for the production¬scale experiment (Figs. 1 to 3). The pillar was about 49 m (160 ft) long, 6 m (20 ft) wide, and 20 to 26 m (65 to 85 ft) high. The mined area on both sides of the pillar was backfilled with a 1:30 cement:sand mixture. The pillar was removed in two phases. In phase 1, the standard uphole method was used to blast down the 18-m (60-ft) long section of waste from the bottom of the ore into the undercut. From the pillar's top sill, 165-mm (61/a-in.) holes were drilled downward into the pillar, and by measuring the depth of the holes, the results of the uphole blast were determined and roof line 1 was established. The bottom of each hole was plugged, then filled with sand to place the center of gravity of each spherical charge (loaded from the top sill) at a predetermined optimum distance from the horizontal free surface. The charges were then detonated. After detonation, both draw drifts were full of extremely well-broken material. The depth of each hole was measured again, and the plot of these depths re¬sulted in roof line 2. The same blasting procedure was repeated and the resultant new back elevation was marked by roof line 3. The poor results of the initial uphole blast at one location (notice the peak in area 1) appeared to influ¬ence the subsequent new backs. A third blast success¬fully evened the back, and resulted in roof line 4. An unblasted slab averaging 6.3 m (20.9 ft) thick remained below the pillar's top elevation as the final sill. In all three spherical charge blasts fragmentation of the blasted material was extremely fine. The backfill was fully exposed on both sides of the now-blasted pillar. The backfill remained undamaged and the ore was not diluted by sand. The remnants of all the 165-mm (61/2 -in.) holes remained clean and undamaged, and the holes had well-defined bottoms that could be easily measured and plugged. Each blast took down a 3.9 to 4.2-m (13 to 14-ft) thick horizontal slab of ore. Productivity was three times greater than that of the previously practiced cut-and-fill method. Since this was the first such experiment, blasting the remaining 6-m (20-ft) thick final slab was the sub¬ject of some deliberation. If the described method was repeated, we could have ended up with a 1.8-m (6-ft) thick sill unsafe to work on. It was therefore decided to blast the whole sill using two spherical charges prop¬erly placed in each hole, but with the application of

Jan 1, 1982

By Frank Florez

INTRODUCTION The Magma mine, located at Superior, AZ, some 100 km (60 miles) east of Phoenix, was discovered in 1875, and was originally mined for silver. The property was acquired by W. B. Thompson who formed the Magma Copper Co. in 1910. Prior to 1965, square-set mining and cut-and-fill methods were used to mine steeply dipping veins at Magma. Ore was moved to the chute by gravity in the square-set stopes and by slusher in the cut-and-fill stopes. The stopes were gobbed with waste from the surface and development headings in the mine. The original square-set method of mining was modi¬fied in order to extract replacement ore bodies in limestones that dipped at 0.52 rad (30°) and varied in thick¬ness from 4.6 to 38 m (15 to 125 ft). Mining of these ore bodies was done by starting at the upper lateral ex¬tremities and retreating to the access openings. The timber square sets and cribbing were used as temporary support, the ground being allowed to cave as the mining retreated. It was found, however, that mining in such a manner was very inefficient; safety hazards increased be¬cause of the blocky and fractured characteristics of the ore and the weakness of the hanging wall. GENERAL DESCRIPTION The undercut-and-fill method of mining is now the primary stoping method used at Magma, and no pillars are left for later extraction. Four ore bodies are pres¬ently being mined. These ore bodies (see Fig. 1) dip conformably at 0.52 rad (30°) to the east with the host limestone and shale sediments, their average strike length is about 122 m (400 ft), and their thickness varies from 9 to 38 m (30 to 125 ft). The ore bodies are separated by varying thicknesses of barren rock. Three of them come together at various elevations primarily in the area of the main feeder fault, but there is an apparent horizontal displacement of about 23 m (75 ft) on the main feeder fault which complicates stope layout. The method as originally applied has been modified in order to mine more efficiently ore bodies that are 9 m (30 ft) or more in thickness. The modification was made to mine all the ore at a given elevation before dropping down for the cut below. The advantage of this is that mining is done on a horizontal plane and there¬fore muck removal can be mechanized. Also multiple faces are made available, thereby improving overall min¬ing efficiency. Another major change made was to carry a 6.7-m (22-ft) face when mining the top sill and a 6.4-m (21-ft) face when mining under fill rather than the former standard 3.35-m (11-ft) high face. This change lengthens the mining cycle and cuts mat prepa¬ration costs in half. Stopes are laid out (see Fig. 2) so that maximum mining panels are available and yet one stope crew will not tie up too much of the ore reserve. One stope might consist of as many as 20 panels, and these panels are divided evenly on opposite sides of the extraction cross¬cut. The stope layout is planned so that the maximum tramming distance to the orepass will not exceed 30.5 m (100 ft) on the crosscut and 38 m (125 ft) in the panel. Stope panel distances are sometimes exceeded when min¬ing to the extremity of the ore body. Tramming distance limits are determined by the efficiency of the com

Jan 1, 1982

By Someswar Kesh, Gavin J. Faulkner

Introduction Powered supports in longwall mines essentially operate under two conditions: 1. Static (non-working) 2. Dynamic (working) There is a variety of static conditions which may exist in a mine. These may include a temporary stoppage of work for a few minutes or stoppage of work for days because of strikes or lack of demand for coal. Studies on pressure behavior were made during brief stoppages of work (few hours) as well as a long stoppage of work (two weeks). In dynamic conditions, the pressure variation on supports is expected to follow a certain pattern. However, in many cases that does not happen. Pressure-time characteristics of each leg of the supports were studied. This paper discusses the research findings of the studies made on powered support pressure behavior under both static and dynamic conditions. Studies on Pressure Variation in Static Conditions It has been noted (Brenner et a1.,1975) that the pressure distribution on mechanized longwall faces is random in nature and changes with time and position along the face. Figure 1 and Figure 2 show the distribution of support pressures for the front left leg of supports in two different mines. In both the cases it can be seen that the variation of the support pressures along the face is indeed absolutely random. However, the nature of randomness in both the cases is very different. The data in Figure 1 (Mine 1) is much more random than the data in Figure 2 (Mine 2). This can also be seen from the standard deviation of the data (81 26 Pa versus 4000 pa) . By comparing the results of ten other sets of data the same conclusions can be arrived at. Since it was found that in Mine 1, the pressure on support legs was extremely random, batches of ten supports each were taken and the mean pressure plotted. This mean pressure line is shown in Figure 1. It was expected that this would smooth out the variation due to individual props and reveal a better picture of the support pressure distribution. For the case cited, the standard deviation is lowered to 5043 Pa from 8126 Pa which indicates that, on an average, the supports tend to take loads more uniformly, however for individual support legs the load on the supports is very random. The significance of this is that, if the supports are selected on the basis of theoretical load calculations alone (based on the roof block concept or some other concept), only the mean value of the support load required will be known. In the case of a pressure distribution with a high standard deviation, some supports selected on the basis of mean value will fail, while others will be supporting low loads. Hence, for proper selection of support capacity, it is not only necessary to know the average load but also the expected nature of the distribution. The only reason why the supports are supporting the roof in Mine 1 well is the fact that the yield load is above the highest level of load attained. If Mine 2, which has a much higher average load, had the same sort of standard deviation, the loads on the individual legs could easily exceed the support load. For example, if the pressure distribution followed a normal curve and the mean pressure was 36,700 Pa with a standard deviation of 9000 Pa, it is possible that 67% of the supports will be within the pressure range of 36,700 plus or minus 9,000 Pa, or 22,700 and 45,700 Pa. Many of the supports could then exceed the yield load of 40,000 Pa. It is clear then that the mean load is an inadequate criterion for estimating support resistance. The nature of the problem is statistical and such an analysis should therefore be made as far as pressure data are concerned. Statistical Analysis of Pressure Data. In general, in longwall mining and other geo-mechanics studies, it is extremely difficult to specify the exact nature of the distribution of a variable. Under these circumstances, probability

Jan 1, 1986

By M. Sears, G. S. Esterhuizen, Z. Khademian

According to the 2010-2019 Mine Safety and Health Administration (MSHA) accident report database, 91% of reported ground control accidents in U.S. longwall mines were caused by roof instability. Gateroads are subject to significant changes in loading conditions from the development to the longwall abutment loading phases. When combined with thinly bedded shale roof, found in many U.S. longwall coal mines, the design of efficient roof support becomes challenging. In previous work, the Bonded Block Method (BBM) modeling of roof by UDEC was validated against field extensometer measurements in a longwall gateroad entry roof at a 180-m depth of cover. The BBM was shown as capable of capturing delamination and buckling of shale roof, one of the main roof instability mechanisms in longwall mines. This paper presents the recent findings on the roof-support interaction using BBM models of the same longwall entry. The effects of cable bolts, roof bolt density, and strap support on potential roof instability are studied. These types of studies are difficult and costly to conduct in an operating mine. Results show how the contribution of bolts, straps, and cable bolts can be quantified using the BBM models. The study demonstrates the potential for BBM numerical models to help understand the complex roof and support system interactions and to assist with optimizing gateroad support systems.

Feb 1, 2023

By Donald G. Dahle

Project Payette is a conceptual nuclear experiment under the Vela Uniform program intended to further evaluate the mechanism of seismic decoupling. Vela Uniform is a Department of Defense program to improve this country's capability to detect, locate and identify underground nuclear detonations. This program evolved from recommendations of the Panel on Seismic Improve¬ment, (Berkner Panel), an officially appointed group of noted U. S. scientists. This Panel, taking note of difficulties in detecting the differences between some underground nuclear detonations and some earthquakes recommended a program of increased basic research in seismology. The theory of decoupling, Which involves possible reducing of seismic motion resulting from underground explosions by reducing the coupling or physical linkage between the explosion and the geologic environment, was investigated by Dr. Albert latter and his associates at RAND during the late 1950's. The concept has been further investigated theoretically by scientists of the Lawrence Radiation Laboratory and the Sandia Laboratory. The theory has been tested experimentally under Vela Uniform Projects Cowboy, Dribble and Sterling. The Salmon Event of Project Dribble was a tamped (with the explosive in firm contact with the geologic formation) nuclear detonation 2700 feet below ground in the Tatum Salt Dome. Project Payette, if found to be feasible and if authorized for actual execution, would become a decoupled experiment in the same environment.

Jan 1, 1968

By Roberto Carranza

The planning of experiments and the importance of statistical interpretation of the results have been recognized for years by scientists and engineers. The Factorial Design of experiments is among the various methods used for this purpose and is being widely applied in all branches of research and production. The basic principles of this tool have been described in detail (1, 2, 3, 4) and several examples of its application in various fields have been published (5). In recent years the mineral industry has also used Factorial Design to improve the efficiency of research, testwork and plant operation (6, 7, 8, 9). This method permits for a qualitative determination of the influence of the variables studied in a process and also a qualitative measure of the interaction of the variables. By graphic means or derivative methods, one may also establish the optimum conditions of the process (ibid 7, 8, 9). It is desirable, however, to develop mathematical methods to evaluate quantitatively these effects and restrict the optimization to the practical limitations of the variables. It is also convenient to adopt a notation system to generalize the Factorial Design.

Jan 1, 1971

By P. Eng. Daniel Kelly

INTRODUCTION The Astoria property is located in the prolifically mineralized Abitibi greenstone belt and is one of the most advanced projects presently under development by Yorbeau Resources Inc. This property, which hosts the Astoria gold deposit, is located in the southern part of Rouyn township, Abitibi area, Quebec, 4 km south of the city of Rouyn-Noranda (Figure 1). The Astoria deposit has been the subject of underground development for over a half century, beginning with Granada Rouyn Mining Co. in the 1930's and followed by Astoria Rouyn Mines Ltd. in the 1940's. Yorbeau Resources Inc. acquired the property in 1984 and has, since then, conducted numerous exploration programs on this gold deposit. The diluted reserves of the Astoria deposit are estimated at 1,3 million metric tons grading 6,14 grams of gold per metric ton (1.4 million st at 0.18 oz/st). This reserve figure is based on the results of surface and underground exploration carried out from January 1986 to June 1987 and includes ore in the proven, probable and possible categories. This paper is divided into two sections. The first part includes a summary of the geological environment including a description of the gold bearing zones and a short discussion on wall rock alteration. The second part of this paper reviews the underground work and the objective of the ongoing exploration program.

Jan 1, 1988