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By G. D. Bates

INTRODUCTION The Magmont mine is a joint venture of Cominco American Inc. (operator) and Dresser Minerals, Inc. The mine-mill operation is located approximately 160 km (100 miles) southwest of St. Louis, MO, in what is commonly referred to as the "Viburnum Trend.” The Magmont mine is designed for a production rate of 3810.2 t/d (4200 stpd) on a 5-day week, three shifts per day basis. Initial production began in 1968. The mine is open stope, room-and-pillar, and essentially horizontal along the trend of the ore body. Briefly, the main geological structure can be described as a brecciated graben bounded by reverse faults. The ore body in cross section is shaped like a bell curve with some lateral extension at the lower part. Presently outlined ore is 609.6 to 762 m (2000 to 2500 ft) in width and 2133.6 m (7000 ft) in length. The ore varies in thickness from 4.87 m (16 ft) on the fringes to an average of 27 m (90 ft) in the high ore areas bounded by the reverse faults. Lead is the primary metal with zinc and copper secondary. MINE DESIGN The basic design of open stope, room-and-pillar mines has been described by several writers and need not be repeated here. (Anon., 1970; Bullock, 1973; Casteel, 1972; Christiansen et a]., 1970; and Lane, 1964) This discussion covers the mining sequence as applied to the particular problems at the Magmont mine, the use of equipment, and deployment of the work force. In the upper portion of the Magmont ore body is a layer locally called the False Davis shale. This layer lies below the true Davis shale, is normally interbedded with dolomite, is of varying thickness, and if mineralized, is included in the top pass of the mining sequence. In thick ore areas this layer will be 2.13 to 2.43 m (7 to 8 ft) in thickness and will occur in the upper portion of the pillars. Due to its incompetency the presence of this False Davis layer is of primary concern in mine planning and operation. Mining areas are divided into three basic groups by ore thickness. First are areas of ore up to 6.09 m (20 ft) in thickness. These areas are below the False Davis shale and are mined single pass with drill jumbo. Second are those areas up 13.71 to 15.24 m (45 to 50 ft) in height. The first 4.87-111 (16-ft) Pass is taken at the top of the ore and the back and pillars secured. Benching the lower portion(s) in 4.57 to 4.87-m (15 to 16-ft) passes is then done with either a drill jumbo drilling horizontally or a crawler drill drilling vertically. Normally these areas are below the Table 1. Productivities per Manshift False Davis shale. These areas may also be mined by back slashing, or overhand benching, where the first 4.87-m (16-ft) pass is taken at the base of the ore and successive 4.87- m (16-ft) passes are taken upward. A minimum amount of back slashing is done at Magmont since it requires repetition of roof control on each pass and roof control is the single largest stoping cost at Magmont. Ore left to provide a working platform oxidizes and is coated by oil spills thus reducing metallurgical recoveries. The third mining area is over 15.24 m (50 ft) in height UP to a maximum of 40.23 m (132 ft) and will encompass the False Davis shale. These areas are mined by first driving +15% inclines to the top of the ore body. The top pass is mined and the back is bolted and roof mats installed as a matter of standard practice to minimize roof problems as mining progresses downward. Once the back and pillars on the top pass are secured, benching begins on successive passes with either the drill jumbo or crawler drill. Pillars on all successive passes below the top pass are secured as necessary. While benching progresses below the top pass, the pass at the base of the ore body is mined leaving a sill of 4.57 to 7.62 m (15 to 25 ft) in thickness to be removed with the crawler drill in a retreating manner. Rooms are mined on a 1.57 rad (90") grid pattern to insure alignment of pillars where multiple passes are taken. Pillars are designed on a 17.98-m (59-ft) spacing with rooms up to 10.66 m (35 ft) in width. Heading widths are wide enough for the mobile equipment to turn without additional allowance for curves. The result is a flexible layout which provides a maximum number of headings available for high extraction rates and grade control. PRODUCTION Incentive Bonus Incentive bonuses play an important part in the mine production at Magmont. Production crews are trained to perform only one of the mining functions of drilling, blasting, mucking. or roof bolting. This specialization, or functionalization, is augmented by development to open all possible stoping areas as early as possible in the life of the mine. This insures that each crew will have enough headings to perform its specialty. The incentive bonuses increase exponentially as output increases. The lucrative incentive bonus coupled with the specialization of the production crews and proper mine development have combined to give the high productivities shown in Table 1. Development crews perform all mining functions in their area. The incentive bonus is paid on a per foot basis, Crews on different shifts working the same heading share equally in the bonus proportional to their contract hours worked.

Jan 1, 1982

By R. Sandefur

Recent advances in personal computer technology allow movies of various types to be made and displayed on PC and Macintosh computers. An interesting application of this movie technology is animation of computerized geological models. These movies, which can be most easily constructed along rows, columns, and benches of block models, allow an ore body to be more easily and quickly visualized than the traditional method of flipping back and forth through various sections.

Jan 1, 1994

By Basar Arioglu, Jaeung Kim, Mehmet Öztürk, Jinmoo Lee, Dongho Lee, Taek-Kon Kim

"Istanbul Strait Road Tube Crossing Project (Eurasia Tunnel) includes 3.34 km of a bored double-deck tunnel under the Bosphorus Strait, at a maximum depth of 106 m from the sea level, and 1 km of twin NATM tunnels on the Asian side. The region’s variable geology, hydrology and propensity for seismic activity, combined with high water pressure and large-diameter/double-deck tunnel configuration, make the Eurasia Tunnel one of today’s most challenging and complex projects under construction. Due to the site’s proximity to the active North Anatolian Fault, the tunnel is designed to withstand a 7.5 magnitude earthquake. The tunnel penetrates through both the Trakya bedrock (on both European and Asian sides) and the alluvial sediments at the bottom of the channel. The tunnel is designed to accommodate differential displacements by introducing seismic/flexible joints at proper locations at the transition zones between rock and soft soils. Tunnel fire is also a significant design consideration for undersea tunnel due to that the rubber protecting it in close proximity of the seismic joint should resist a fire for long term safety and durability. This paper mainly focuses on the design for method of tunnel response analysis in transverse and longitudinal direction against seismic demands relating with the seismic joints. INTRODUCTION Istanbul Strait Road Tube Crossing Project (Eurasia Tunnel) will build and operate a roadway tunnel under the Borphorus connecting Kazlicesme on the European side and Goztepe on the Asian side, to provide additional road-access across the Strait and consist of three main parts as shown in Figure 1."

Jan 1, 2016

By M. I. Brittan

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

Jan 1, 2010

By D. Nebbia, R. Aliberti, F. Solesin

The launch of a Tunnel Boring Machine (TBM) is one of the most challenging activity parts of the construction of a tunnel. Challenges are coming from limited space available, learning curve, concurrence of other activities and in some cases from the geological conditions. Launching a TBM is an extraordinary and complicated operation requiring a great level of planning. The Earth Pressure Balance (EPB) TBM named “Chris” was successfully launched for the Northeast Boundary Tunnel, the largest component of the Clean Rivers Project in Washington, DC. Due to space and time constraints, the 8-meter diameter TBM was launched from a single 20-meter diameter shaft, with a 30-meter-long tail tunnel by means of umbilical connections. This paper presents the steps followed during the TBM assembly and launching to complete the excavation of the initial drive of the tunnel maintaining high level of efficiency and quality despite the space constraints. Specifically, it focuses on the methods and phases used to extract the excavated material (muck) from the TBM and from the tunnel, to supply consumables, precast rings and backfill grout. The paper presents results in terms of productivity, lesson learned and possible alternatives to be considered in similar conditions.

Jun 13, 2021

By Rakesh Sehgal, Pushkar Verma, Mohit Sharma

"SJVN Limited successfully commissioned its Rampur HEP (412 MW) in 2014. The project has been planned to operate in tandem by directly utilizing the Tail Water of SJVN’s prestigious Nathpa Jhakri Hydro Power Station 1500 MW located upstream. The initial filling of voluminous water conductor system of RHEP with 10.50 m diameter, 15 Kilometer long Head Race Tunnel, involved more than 1.5 million m3 of water. The filling operation of such a large diameter tunnel constructed in young folded Himalayas which always offer less ideal sites for tunnelling and their subsequent operations under pressurized conditions is worth sharing with tunnelling fraternity. The paper covers meticulous planning, execution, precautions taken, and performance monitoring during filling process keeping in view the fragile geology, low rock cover zones, shear zones in close vicinity of thickly inhabited areas, squeezing zones, hot water zones encountered during excavation.1.0 RAMPUR HYDRO POWER STATION IN BRIEFRampur Hydroelectric Project (412 MW), is located in Himachal Pradesh (India). This project is a downstream development of Nathpa Jhakri Hydro Power Station (1500 MW) constructed on river Satluj and commissioned during 2003. Both the projects today are now successfully working in tandem. The upstream NJHPS has been constructed on the left bank of the river while the downstream Rampur HPS is located on right bank of the river (Figure-1).The major length of HRT of Rampur project lies on right bank of river before it crosses under the river in its initial reach (Figure-2). The main features of the project are as below:• 10.50 m diameter 15 km long HRT ending in a 155.75 m deep, 38.00 m diameter surge shaft.• 3 steel lined pressure shafts, 5.4 m dia, 211-220m long.• Surface valve house 69 m long x 10.50 m wide x 23 m high housing 3 No. butterfly valves• 3 No. surface penstocks 5.4m dia 368-382m long, bifurcating into six penstocks, each of 3.80 mdia. to feed six generating units of surface power house 158 m long x 24.50 m wide x 48 m high, with Francis turbines each of 68.67 MW capacity."

Jan 1, 2016

By Don Painter, Nico Sutmoller, Lori M. K. Kahikina, Brian Dorwart

This paper will address the feasibility of using low-density cellular concrete (LDCC) for an annular fill application where project conditions required pumping the LDCC material for distances of up to 16,000 feet. These conditions occurred during construction of the Kaneohe-Kailua Sewer Tunnel in Oahu, Hawaii, between 2014 and 2018. Located on the island of Oahu, Hawaii, this 13-foot-diameter rock tunnel extends for approximately three miles from the Kaneohe Pre-Treatment Facility to the Kailua Regional Wastewater Treatment Plant. The tunnel was built to convey wastewater and provide temporary storage for peak wet weather flows entering the collection system, thereby reducing the potential for uncontrolled discharges. LDCC was chosen for this project partly due to its environmental benefits, which were essential for an application contiguous to the inland marine and coastal waters of the Pacific Ocean. In order to ensure optimal performance characteristics—compressive strength (with a requirement of 300 psi at 56 days), flowability, and light weight—were achieved, extensive testing of various LDCC products and mix designs was completed. The testing measured the initial set time of the LDCC as well as the compressive strength needed to reduce the quantity of grout lifts required for the project.

Jun 13, 2021

By J. G. Pippenger

Rio Tinto Coal Australia (RTCA) developed Meandu Mine in Queensland, Australia in 1983 to supply thermal coal to the adjacent newly constructed Tarong Power Station (TPS). In designing the mine, RTCA selected Meandu?s best economic reserves for a 25-year life based on station owner Tarong Energy Corporation?s (TEC?s) estimated life for TPS. That 25-year plan changed dramatically in the early 2000s when TEC approved life extensions to TPS?s four existing units plus construction of a fifth unit. After reviewing Meandu?s remaining reserves, RTCA decided the nearby Kunioon reserve offered the more economic option for maintaining deliveries beyond 2010. Those plans changed yet again in 2007 when TEC chose to purchase the Meandu Mine and nearby reserves rather than extend RTCA?s contract. TEC then selected Thiess Pty Ltd (Thiess) to continue operating Meandu until the new mine at Kunioon could be developed.

Jan 1, 2011

By J. Sottile, V. Androulakis

Operators of underground coal mine shuttle cars are exposed to numerous hazards including noise, dust and musculoskeletal trauma. The repetitive nature of shuttle car operation leads to fatigue-related incidents and soft tissue injuries. The poor visibility conditions, and dust and noise increase the occupational hazards to the operator, as well to other miners. To reduce the risks for all miners, towards providing a safer working environment for all workers, the introduction of autonomous shuttle cars is essential. This paper introduces a project to demonstrate the feasibility of incorporating autonomous shuttle cars in the underground coal mining cycle. The autonomous shuttle car will be able to localize itself, to map its surroundings and to navigate in an underground GPS-denied environment. It will also navigate efficiently by fusing different on-board sensor modalities, autonomously planning the optimum path and traversing around the mine while avoiding collisions with humans and other static and dynamic obstacles. This will be achieved by exploring and testing different state-of-the-art techniques and off-the-shelf components, and integrating them into existing equipment, i.e., retrofitting an actual shuttle car. Moreover, a modified mining system will be demonstrated which will incorporate the autonomous shuttle car, capable of making real-time decisions, into the mining cycle. The system will regulate the interactions between the autonomous machinery and the human dynamic, with emphasis given to the safety of the workers, and enable the real-time, remote supervision of the mining process. The focus will be on coal mines, where the room and pillar mining method is used. INTRODUCTION The growing demand for mineral supplies combined with the steady decline in the available surface reserves has driven the mining industry to mine deeper underground deposits. These deposits are difficult to access, and the environment may be hazardous to mine personnel (e.g., increased heat, difficult ventilation conditions, etc.). Moreover, current mining methods expose the miners to numerous occupational hazards such as working in the proximity of heavy mining equipment, possible roof falls, as well as noise and dust. In recent decades, the utilization of autonomous machines has evolved in the mining industry to provide increased health and safety to mine personnel and reduce their exposure to adverse conditions. Automation aims to remove the workers from the hazardous conditions which may be present in an active mining area and place them to a more comfortable and safer environment at a distance from the active mining face. Examples of incorporation of autonomous machinery in the mining industry include the fleet of more than 100 autonomous mine trucks that are employed at the Rio Tinto’s iron ore mines in Western Australia for material haulage. There are also semi-autonomous, teleoperated LHD machines employed at several mines around the world (e.g., Sweden, USA, China). However, the operation of the autonomous machinery in these cases is either partial (semi-autonomous, teleoperated vehicles) and of small-scale or occurs at surface mines where GPS significantly facilitates the autonomous navigation of them.

Jan 1, 2019

By R. Thorpe, M. Dosbaba, A. Kopriva, A. Niiranen, J. Ruokanen

"Automated Mineralogy (AM) is an important diagnostic tool in flotation plants as it adds ore and particle mineralogical properties to traditional mass and chemical data. AM is used in project development, for future ore studies and for diagnosing concentrator operations. With the exception of operations who own central laboratory or on-site systems, the use of AM is often limited to the analysis of a comparatively small set of samples. Metrix Plant Technologies has partnered with TESCAN to integrate TIMA AM technology into the implementation of Performance Assessment Campaigns (PAC). Routine daily composite samples from the mill are analysed mineralogically on a size-by-size basis over several weeks. All the relevant metallurgical, process control and mineralogical data is combined into a single large database for mass balancing and interactive examination. This paper describes the implementation and results of a PAC at the Agnico Eagle Kittilä gold mine in Finland. Although challenging, the addition of mineralogical data as dependent variables in multivariate analysis leads to a much deeper understanding of the root cause effect of events driving the flotation performance. INTRODUCTION TO STUDY The Kittilä mine, owned by Agnico Eagle, is Europe’s largest primary gold mine, with annual production of approximately 180,000 Oz. The operation is located in Lapland in northern Finland, and is currently expected to operate until 2035. The location of the operation is shown in Figure 1. Historically, the mine was an open pit, although in recent years ore is sourced solely from underground operations. The processing plant features crushing and grinding, followed by carbon flotation and sulphide flotation. Carbon concentrate is a waste stream, and sulphide concentrate reports to pressure oxidation (autoclave) and CIL circuits for gold extraction. The final flowsheet is shown in Figure 2. Following commissioning in late 2008, the operation achieved commercial production on May 1st 2009. Gold recovery in flotation gradually increased to a 92-96% range at a throughput of ~3500-4000 tonnes per day. In late 2014, the operation commissioned a plant expansion to ~5000 tonnes per day. Following a commissioning period, the operation stabilised with recoveries in a 90-94% range, which over any given period was determined to be ~2% lower with statistical significance. The trends from start-up of gold recovery in flotation and throughput are shown as follows in Figure 3. In early 2016, Agnico Eagle Kittilä mine, along with Agnico Eagle Technical Services (AETS) partnered with Metrix Plant Technologies and TESCAN, to perform a Performance Assessment Campaign (PAC) at Kittilä, involving automated mineralogy on five key streams at Kittilä on 40 days selected from a 50 day period between February 15th and April 4th 2016, when plant operations were stable. In addition to the mineralogical data, metallurgical and operational data were pooled into a single, powerful database to explore opportunities for the operation to improve recovery in a sustained manner, ideally to the levels pre-expansion."

Jan 1, 2017

By Dapeng JIANG, Haixiao YANG, Shanxi WU, Jinwang Zhang, Guoliang XU, Yipeng ZHANG

The subsidence and deformation characteristics of strata with thick unconsolidated overlying strata is quite different from that with consolidated overlying strata. In this paper, based on the engineering background of large mining height mining in No.21 coal seam of Zhaogu No.2 Coal Mine in Henan Province, China, the PFC 2D numerical simulation and surface subsidence field observation was conducted to study the subsidence characteristics of strata with thick unconsolidated overlying strata (680m). Numerical model with dimensions of 2000m 750m was established and mining process with different working face lengths (150 m,200 m,250 m,300 m) and different working face advancing distances (40 m,80 m,120 m,160 m,200 m,240 m,280 m,320 m,360 m,400 m,440 m,480 m) were simulated. The surface subsidence characteristics and displacement field variation of rock strata under different working face advancing distances were observed and analyzed. The results show that with the advance of working face, the structure of subsidence-like basin will be formed in each rock stratum, and the influence range of mining on rock stratum is “bowl-shaped”, and the influence increases with the increase of advancing distance. The unconsolidated layer has a buffer effect on the mining-induced subsidence, and the buffer effect is realized by the expansion of the subsidence area. The movement law of strata under different working face lengths is similar to that under general mining, and the structure of subsidence-like basins will also be formed in each stratum. The influence range of strata after mining is ' bowl-shaped ', and the influence extent increases with the increasing of working face length; the internal subsidence of the unconsolidated layer is more intense and complex than that of the rock layer; the surface subsidence movement shows an obvious 'basin' phenomenon. The field observation data of surface subsidence also confirm the above results.

By G. J. Joy, J. P. Rider

"Technical advances in longwall mining over the last several years have resulted in much larger and faster shearers that have the capability of mining at speeds over 100 ft./min. This has resulted in a substantial increase in the amount of dust that is generated at the shearer and during shield advances. Meanwhile, increased underground coal production has put a greater demand on longwall dust control systems as some operators have had difficulty maintaining consistent compliance with federal dust standards. To address this need, the National Institute for Occupational Safety and Health [NIOSH] is investigating the effectiveness of a water spray manifold mounted on the tailgate end of the shearer body. The sprays are oriented parallel to the tailgate ranging arm, and these sprays create a water curtain confining the dust plume near the face, preventing the dust from drifting into the walkway. This redirection of dust provides a clean air envelope for the tailgate shearer operator and jacksetters working near the tailgate drum. Three different tailgate spray manifold designs were tested and evaluated at the Office of Mine Safety and Health Research [OMSHR] Pittsburgh Longwall Test Facility. Manifolds were equipped with hollow cone or flat fan water spray nozzles and evaluated at three face air velocities (500, 700, and 900 fpm) and at water pressures of 100, 150, and 200 psi. Reductions in dust concentrations were observed at six face sampling locations for all the tailgate manifold systems at each of the tested spray pressures and face velocities. Based on results from the laboratory tests showing reductions in dust concentrations at sampling locations downwind of the shearer a spray manifold was fabricated and evaluated at an underground operation. INTRODUCTION In 2012, underground coal mining in the United States produced a total of 342 million short tons, with longwall mining accounting for over 53% of this production [EIA 2013]. The longwall mining industry continues to see remarkable and significant improvements in longwall mining equipment and mining practices. Average shift production during compliance dust sampling has increased from 3,500 tons per shift in 1994 to approximately 6,000 tons per shift in 2012. A dramatic decrease in working longwall faces from 80 to 49 has occurred over the same period [Fiscor 2013]. Factors specific to longwall mines that supported productivity improvement include: higher capacity shearers, improved shield hydraulics and control, higher horsepower for equipment, wider face conveyors, and larger faces. Today, the average face width has increased to 1,188 ft., with one longwall operation reporting a face width of 1,650 ft. compared to an average of 750 ft. in 1994 [Fiscor 2009, 2014]. Panel lengths in 2013 averaged 11,307 ft. compared to 7,000 ft. in 1994. Overall production from U.S. longwalls has increased by nearly four percent from 2007 to 2012 with over 182 million tons mined annually [EIA 2013]. Given the volume and rate of coal produced on longwall faces, the potential to generate large quantities of respirable coal mine dust also exists."

Jan 1, 2015

By H. Weber

This paper is in three parts. Part I discusses the various types of cement clinker coolers now in full-scale operation. Part II is a detailed study of the criteria for selection of the type of cement clinker cooler best suited for a given installation. Part III covers a number of cooling problems and their solutions in some of the metals industries. The large and world-wide problem of cement clinker cooling has motivated much study and development and the coverage here of this work is somewhat detailed. In contrast the cooling problems of the metals industries are quite varied between industries and often quite different from plant to plant in the same industry because of basic differences in the requirements of the particular material being processed. Therefore the coverage of metals industries coolers is somewhat more general. It is believed that all of the technology covered here may be of interest to those working with the problems of cooling hot materials.

Jan 1, 1980

By Frederick McNeill, Mahmood Khwaja, Benjamin Lundsted, David Polcari, Michael S. Schultz, Timothy Clougherty

The City of Manchester, New Hampshire, is embarking on one of its largest public works projects in history to construct a 12,000 foot long, large-diameter, conveyance tunnel through downtown Manchester. This area has challenging geology with undulating bedrock (granite, schist, and gneiss) profile overlain by fluvial deposits. Understanding the subsurface conditions is key to mitigating project risks, establishing plan and profile for the tunnel, and selecting the most cost-sensitive and technically appropriate tunneling method. This paper presents the project team’s approach for conducting the ground investigation, developing preliminary subsurface profile, and the considerations for selecting the tunneling method.

Jun 13, 2023

By Brain W. Lawrence

Sublevel stoping generally is a large-scale open stoping method. It sometimes is referred to as long-hole or blasthole stoping. This method usually is applied to strong ore bodies that require minimal support and are surrounded by strong country rock. The ore body should be fairly regular in shape and well-defined. The dip of the footwall normally should be sufficient to allow broken ore to gravitate freely, although the method has been adapted successfully to mine some of the flatter ore bodies. Sublevel stoping is not dependent on the width of the ore body. Widths of less than 6 m (20 ft), however, make the utilization of long-hole drilling techniques more difficult. The main criteria for sublevel open stoping are competent ore and stable host rock, regular ore boundaries, and a footwall dip that exceeds the angle of repose of broken ore. Basically, the method entails providing access to the ore body at various subintervals between the main haul- age levels in order to drill and blast the intervening ore. Stope drilling is carried out from drilling drifts on the sublevels, and the ore is blasted in slices towards an open face, which generally is vertical on the downholes and may be inclined towards the open face for the up holes. The blasted ore gravitates to the bottom of the stope and is collqcted through drawpoints. Fig. 1 illustrates a typical sublevel stoping mine. The method requires extensive ore body development with relatively high capital expenditures. How- ever, much of the development is in ore, and production costs are comparatively low. Productivity rates are in the 13.6 to 27.2 t (15 to 30 st)/manshift range. The drilling, blasting, and loading operations are performed independently, and equipment utilization is high. Large outputs can be obtained with few units and limited personnel. Dilution with waste rock may occur if ore boundaries are irregular or if caving occurs, but 100% of the ore within the stope usually is recovered. Pillar recovery sometimes is a problem. It is at this point that most of the waste dilution occurs. At the time data was gathered for this report (1973), the sublevel open stoping method was used by more than 21% of the metallic ore mines in the US (9 out of 42 mines producing 1088 + t/day (1200 + stpd) and provided more than 12% of the total metallic ore production from these mines. DEVELOPMENT Access to a sublevel open stope mine is gained by a slope or shaft that is normally sunk in the footwall of the ore zone, away from any possible effects from blasting or other production operations in the stoping process. Main level interval is selected and usually varies from 45.7 to 121.9 m (150 to 400 ft), depending on the vertical extent of the ore body, the number of drawpoints required to maintain output, and the eventual stope height. The main haulage drifts are located at the bottom of the stope horizon, either directly in ore or in the footwall, with access crosscuts at intervals. Raises are driven up through the ore to connect to the level above and provide access and ventilation to the sublevel drilling drifts, which are driven horizontally in the ore for the length of the stope. Depending on the drawpoint system to be used, the stope is either undercut by a raise-and-cone system, from a scram or slusher drift, or simply provided with a drilling drift from which access is made at intervals to the main haulage level. A raise driven either at the end or in the middle of the stope (if stope access is from both ends) provides a start for opening up a vertical slot across the stope to establish a free or open face for long-hole drilling and blasting. Sublevel stoping has proved to be a relatively safe and economical system.

Jan 1, 1982

By Allan J. Breitenbach

Heap leach operations provide a low-cost method for recovering gold, silver, and copper metals from low grade ore materials. The processed ore materials are stacked in controlled lifts on a lined pad facility for heap leaching. A controlled barren solution is applied to the heap ore surface by sprinkler or drip emitter irrigation, and ore metals are leached into the solution for collection and extraction in plant process facilities. The leach pad and solution collection facilities are lined with impervious materials for containment of all process solutions. The liner system typically includes a composite geomembrane and clayey soil liner with a protective drain cover fill to maximize solution recovery and protect the underlying groundwater conditions. This paper discusses the types of pad liner systems currently used by the mining industry and the "good, bad, and ugly" lessons learned in engineering practice over the last 10 years in the design and construction of heap leach pads.

Jan 1, 1999

By M. Nambiar

Due to the remote locations of the mining projects and the demand for long transmission lines to provide electrical power, mining sites often operate with poor network power quality and low short circuit capacity. In order to cope up with this situation, more and more mining projects are considering the installation of appropriately designed Synchronous Condensers(SC) together with Power Factor Correction and Harmonic Filter System(PFC&HFS). The installation of these power quality equipments is viewed as a technological advancement in the mining application. An adequate level of short circuit power is essential and has to be ensured for the safe and reliable operation of especially large cycloconverter driven gearless mill drive systems used for grinding application. Increase in plant short circuit power is generally achieved by lowering the impedance of the supplying network. The SC also provides smooth voltage control during mill starting or major load trips, thereby regulating the plant voltage within the required limits. This paper will explore the impact, technical advantage and reasoning for the selection of the advanced hybrid technology mentioned above, in order to cope with the power system challenges for future mining projects. It also highlights the potential problems and how to address them carefully.

Feb 23, 2014

By H. O. Harju

In July 1968, P.T. Inco Indonesia was granted a Contract of Work to explore and develop the nickel laterite resources on the southeastern peninsula of the island of Sulawesi. Since 1970, exploration has centred on the deposits at Soroako as the site for initial project development. This paper describes the exploration methods used to evaluate these rather unique laterite deposits, the geology of which is described in the first section of this volume.

Jan 1, 1979

By Donna L. Boreck

Introduction Coal has been mined in the western United States since the mid-1800s. The resource is abundant in the West. Despite recent economic conditions, it remains an important part of the nation's energy warehouse. Many of the seams that comprise the western resource base are thick, some exceeding 61 m (200 ft). Multiple, closely-spaced seams or thick coals that split into two or more minable seams are also common. Historically, subsurface development recovered from less than 15% to 50% of these deposits (Matson and White, 1975). This resulted in a substantial loss of inplace reserve due to the limited height that could be mined using the available technology. In an effort to increase future recovery from such deposits, the US Bureau of Mines analyzed the potential for incorporating three thick-seam mining methods into western coal mines. These methods are high-face longwall, multislice longwall, and longwall caving. They were developed in Europe and are being used to mine thick coal seams in Europe and Asia. One aspect that will determine their success in the US is the geology of the individual deposit; namely, are the methods adaptable to the geologic conditions characteristic of western coal? This paper analyzes some of the geologic factors that would affect development of thick or closely-spaced seams using these mining methods. Occurrence of thick and closely-spaced seams Since mining conditions and methods differ around the world, a thick seam is often defined as a seam whose total thickness cannot be extracted using the available equipment. In the western US, the maximum extractable mining height is about 4.3 m (14 ft). For this research, a thick seam was defined as a minable seam 4.6 m (15 ft) thick or more. Since one of the seams is usually lost using standard mining methods, closely-spaced or split seams were also addressed. These were defined as two minable seams 1.5 m (5 ft) thick or more separated by 9 m (30 ft) of interburden. Thick or closely-spaced coal seams are common in the West. Their importance has not been documented due, in part, to the methods used to measure resources. In the past, all reserves greater than 1.1 m (42 in.) thick for bituminous seams and greater than 3 m (10 ft) thick for subbituminous seams were grouped together and were differentiated only by rank. Further differentiation by seam thickness or minability was not done. For this reason, reserve figures for thick or split seams are difficult to determine. By analyzing available published resource data, Ferm and Muthig (1982) addressed thick seam resources. They estimated a resource of 90 Gt (99 billion st) of coal (16% of the Western resource base) in seams 4.6 (15 ft) or greater at less than 610 m (2000 ft) overburden. Yet, a later study of Pierce and others (1982) on one thick seam deposit in the Powder River Basin estimated a resource of 103 Gt (113 billion st). Pierce (1986) reported that the overburden reached a maximum of 731 m (2400 ft). The resources in this single deposit exceeded the 90 Gt (99 billion st) total projected for the entire western US by 12.7 Gt (14 billion st). Due to such differences, this author chose to look at reported occurrences of thick or closely-spaced minable seams in place of resource figures, to verify the need to research thick seam mining in the West. An occurrence was defined as an isolated site where a thick seam or two or more closely-spaced seams were reported. The occurrences were derived from published material - mostly mine descriptions, measured sections, coal exploration logs, and oil and gas logs. Due to time constraints, the reserves associated with each occurrence were not determined. All data points were random. The analysis was limited to occurrences in Colorado, Wyoming, and Utah. Although source material available to the author was limited to general coal summa-

Jan 3, 1988

By E. G. Hoover

In the mid-I 970s the rallying cry sent out by the aggregate trade associations to aggregate operators was to come to Washington and help design national legislation mandating reclamation of pits and quarries. This collective effort paid off in that aggregate operators were able to assist in the design of "livable" legislation; however, the coal mining industry took some very hard hits. A principal argument used by the aggregate industry was that relatively small parcels of land are impacted, the operations are located in or near urban areas to serve in the development of that immediate area, and are of many years duration; whereas, coal strip mines disturb many acres, are of short duration, and whose products may have world or nationwide utilization. Historically, developments in reclamation are tied closely to the regulatory systems that have been created to control environmental effects of mining. Since reclamation generally results in a cost incurred after the stone, sand and gravel has been mined and does not add to the inherent value of the mineral, there was little incentive in the past for mine operators to reclaim the land. In 1977, the Surface Mining Control and Reclamation Act (SMCRA) became the law for coal mine operators. The law states that, among other provisions, reclamation must "restore the land affected to a condition capable of supporting the uses which it was capable of supporting prior to any mining, or higher or better uses". It is of interest to note that years before the initiation of legislation leading to SMCRA twenty-eight states already had reclamation regulations that covered virtually all minerals. The scope of performance standards found in the state reclamation regulations included slope stabilization, grading, revegetation, neutralization of toxic materials, visual screening, waste disposal, control of drainage, and other sources of pollution and erosion. Environmental legislation and the consequent regulations came about because the public perceived that private industry did not properly value environmental quality and as a result the government became the policeman.

Jan 1, 1998