Search Documents

"The Ernest Henry Mining (EHM) ore body is an iron oxide copper gold deposit with an average grade of 1.1 percent copper, 0.5 g/tonne gold and 22 percent magnetite. At the end of 2009, EHM converted the ball mill media charge from forged steel to high chrome media. At high pH values (above 10.5), forged media is passivated, so as the pH of the system drops the media corrosion rate increases and floods the system with corrosion products, affecting flotation. As the high chrome media is much more inert, pH has less of an effect on copper flotation response. Given this, the pH in the roughers was dropped from 10.6 to 9.6 and it was found that there was a slight increase in copper recovery and no statistical change in either the rougher or final copper concentrate grades for copper and gold. With the reduction in pH, there was a noticeable reduction in lime usage from 0.85 kg of lime per tonne of ore processed to 0.55 kg/t. This reduction in lime is equivalent to approximately $700,000 per annum at current production rates.INTRODUCTION The Xstrata Copper Ernest Henry Mine (EHM) is situated 38 km north east of Cloncurry, Australia. The mine began as an open-cut operation in 1998 and transitioned to a multi-sourced operation in December of 2011, utilising ore from the newly completed underground mine in addition to ore mined from the Mount Margaret Mining operations. EHM produces a combined copper gold concentrate containing 25 000 t of copper and 35 000 oz of gold annually, utilising a three stage grinding circuit along with a four stage flotation circuit. In addition to this EHM also produces 1 000 000 t of magnetite concentrate annually via a further grinding and three stage magnetic separation circuit. Based on metallurgical test work performed by Magotteaux between 2002 and 2008, the step to change from forged steel media to 18 percent chrome media was taken in 2009 in a bid to alter the pulp chemistry to improve the flotation performance in the plant (Greet, Kinal, Twomey, & Peatey, 2010). This plant trial resulted in: • Increased dissolved oxygen levels within the pulp, • A ball mill discharge Eh shift of 25 mV to more oxidising pulp potentials, and • An EDTA extractable iron percentage reduction. These improvements in the pulp chemistry resulted in an increase of 1.7 ± 0.6 percent in copper recovery and an increase of 2.9 ± 1.7 percent in gold recovery. As the plant trial using high chrome media was successful, the use of the media was adopted as the standard for the ball mill. This implementation opened the door for further process improvements around the flotation circuit."

Jan 1, 2014

By J. He

The separation performance of a raw coal sample from South Africa based on the dense gas-solid fluidized bed beneficiation technique was investigated. Analysis showed that the raw coal sample distributes mainly in the -50+6 mm size range with the -1.6 g/cm3 density fraction accounting for 78.20% of the cumulative mass distribution. A flow sheet of the sequential coal beneficiation technique was introduced. The experimental and simulation results of the hydromechanical and density distribution stability indicate that the dense gas-solid fluidized bed could provide a stable separation environment with uniform density fluctuations under suitable operating conditions. With the application of this technique on the raw coal sample, the separation results show that three products, consisting of 73.25% clean coal with an ash content of 9.59%, 11.31% middlings, and 15.44% gangue were effectively separated at two different densities of 1.82 g/cm3 and 1.57 g/cm3. This study provides an effective flow sheet for achieving the high-efficiency separation performance on raw coal from arid areas. Keywords dense gas-solid fluidized bed, dry coal beneficiation, density distribution, separation efficiency.

Jul 1, 2013

Fines and unnecessarily small rock fragments are created by as many as four identifiable breakage mechanisms which operate close to the charged section or sections of the blasthole. The peak blasthole pressure (Pb) should not exceed that value at which the rock starts to fail in compression. Excessive Pb values can be prevented by carefully selecting:1. the composition, density and/or diameter of the charge, and2. the initiation/priming system.The percentage of fines increases with:1. increases in the charge diameter:blasthole diameter ratio, burden distance, blasthole diameter, effective sub-drilling and size of blast, and2. decreases in blasthole spacing, stemming height and the length:width ratio of the blast block.Decked charges tend to reduce the percentage of fines.Excessive fragmentation is more evident in buffer blasting and sinking cut shots than in blasts fired to a clean free face. Initiation sequences which give effective burdens considerably less than drilled burdens encourage the minimization of fines, as does an inter-row delay time which promotes good progressive relief of burden.The wastage of explosive energy in creating fines is most pronounced where fully-coupled, energetic charges are fired in weak and/or highly porous rocks. This is especially so where forward movement of the strata is impeded by buffer rock, excessive burden distances and/or inadequate inter-row delay times.

Jan 1, 1979

By S. Roy

Soft laminated iron ore (SLO) having a higher hematite content and Martite goethite iron ore (MGO) having a higher goethite content were subjected to detailed characterization and beneficiation. The present research was designed to investigate if these abundant, low-grade ores can be enriched adequately using standard industry techniques. No significant enrichment of the ores using jigging operation on the coarse (-3,200 +595 mm) size fraction could be achieved. Good liberation was only attained in the 150 to 300 mm particle size range in both cases. Desliming enriches the ground ores significantly. The effect of the feed size on the separation efficiencies of the ground ores were studied using flowing-film concentration on a Wilfley table. Better enrichment is obtained in the case of 150-mm sized feed by tabling. The grade of SLO is improved from 56.14% to 66.51% Fe, while the Fe content of MGO is enriched from 55.16% to 65.83%. Substantial rejection of alumina and silica is also achieved. Due to the friable nature of MGO and the higher goethite content in it, the loss of Fe in the form of ultrafines is greater in this case compared to that of SLO.

Jan 1, 2009

By Andrew E. Nevin

The government of India convened a precedent-setting meeting in New Delhi on Nov. 12, 2001.The meeting brought together regulators and the mining industry in an attempt to fix the industry’s mine development problems in that country. More than 150people attended the all-day session. Ram Vilas Paswan, Minister of Coal and Mines, chaired the meeting. Ravi Shankar Prasad, Minister of State for Coal and Mines, was also present. The secretary of the Department of Mines and the Joint Secretaries and Additional Secretary also participated. The Department of Mines’ mandate is for economic growth: to get new mine development going and to make investment in mining attractive for both domestic and foreign-owned companies.

Jan 1, 2002

By Raymond W. Henn

The 32nd annual short course ?Grouting Fundamentals and Current Practice? was held at the Colorado School of Mines, June 13-17, 2011. The field demonstration portion of the course was conducted June 16 at Baski Inc.?s yard in Denver, CO. The full-scale field demonstration presents many types of drilling and grouting equipment in operation as well as numerous grouting methods per-formed under various field conditions. As part of the field demonstration, the class was shown the proportioning, mixing, testing and injection of various cement grout mixes into sand columns under controlled and recorded conditions. These sand column demonstrations have been conducted under controlled and measured conditions each year since 1999 as part of the short course. The sand column demonstrations were conducted prior to 1999, but with less quality control and minimal record keeping of the proportioning, mixing, testing, injection pressures and the final permeation results.

Jan 1, 2011

By Thomas G. Goonan

Copper production through smelting generates large volumes of material flows. As copper contained in ore becomes copper contained in concentrate to be fed into the smelting process, it leaves behind an altered landscape, sometimes mine waste, and always mill tailings. Copper concentrate, fluxing materials, fuels, oxygen, recyclables, scrap and water are in-puts to the process. Dust (recycled),gases — containing carbon dioxide(CO2) (dissipated) and sulfur dioxide (SO2) (mostly collected, transformed and sold) and slag (discarded or sold) — are among the significant process outputs. This article reports estimates of the flows of these input/output materials for a particular set of smelters studied in some countries. They include Chile (Altonorte, Caletones, Chagres, Chuquicamata, LasVentanas and Potrerillos), Japan(Naoshima, Onahama, Saganoseki,Tamano and Toyo), the United States(Garfield — Fig. 1 — Miami and Hayden) and Europe (Glogow I, Glogow II, Huelva, Norddeutscheand Ronnskar).

Jan 1, 2005

By M. Kawakami, K. Suzuki, T. Takenaka, S. Yokoyama

Molten NaNO3-NaNO2-KNO3mixture is used as a heat transfer medium at high temperature. One of the most important properties of this mixture is the heat capacity. In the present work, the enthalpy change with temperature was measured by the drop calorimetric method to evaluate the heat capacity. The results are summarized as follows: The enthalpy of the mixture increased linearly with the temperature. From the slope of the line, the heat capacity was evaluated. The heat capacity of pure NaNO3, NaNO2and KNO3was 0.129,0.110 and 0.142 kJ/mol/K, respectively. These values are in fairly good agreement with the published data. In the binary and ternary systems, the molar additivity of heat capacity, which is well known as Neumann-Kopp’s law, was obtained. The heat capacity of HTS1and HTS2, which are well known ternary mixtures of practical use, were in the range of published data.

Jan 1, 2004

ALTHOUGH the development of the various flotation processes during the last 8 years has been so revolutionary, this development has been, until recently, entirely confined to the separation of the gangue from the mixed metalliferous sulphides.The desirability of finding some method of separating the independent sulphides from such a miXture, by means of flotation, has long been recognized. ThIS need was emphasized when the slimes resulting from the grinding of Broken Hill sulphide ores were for the first time successfully dealt with by the Mineral Separation Company's process on the Central mine. This process proved the practicability of separating the sulphides of lead and zinc as a mixed concentrate from the gangue present in slime. Although this was a very notable achievement, the process has not been generally used, owing to the mixed lead and zinc concentrates resulting from the treatment being acceptable neither to the zinc nor the lead smelters.Although the introduction of eucalyptus oil in minute quantities as an emulsifying agent has simplified the work of the subsequent tabling of mixed sulphide concentrates, rendering unnecessary the burning off or removal of the oil, there will always be a considerable quantity of slimy product too fine to settle on the moving surface of the concentrating machines, and therefore requiring other means for its division into lead and zinc concentrates.

Jan 1, 1912

By McHenry Mosier

The Bureau of Mines, United States Department of the Interior, has established a testing adit on the Government reservation at Mount Weather, Virginia, to facilitate research on the fundamentals of mining under actual metal-mining conditions. A number of sites in several Eastern States were investigated, but Mount Weather was selected because the rock seemed best suited for rock drilling and other mining tests and because the ground was already owned by the Federal Government, the site having been acquired in 1902 for use of the Weather Bureau. The testing adit has an advantage over an active metal mine for certain investigations because variable conditions affecting any particular research problem can be controlled and maintained in a fixed status, whereas in an operating mine this seldom is feasible. Operations were begin in September 1936, with a crew of four laborers under a resident engineer. The laborers were recruited from nearby farms, and although they had no previous mining experience, they have developed into satisfactory miners.

Jan 1, 1939

By Greg Rasmussen

"Huckleberry Mines Ltd is an open pit porphyry copper-molybdenum mining operation located 120 km south of Houston, British Columbia with a 18,000 tpd milling facility. Huckleberry Mines is owned 60% by Princeton Mining Corporation, the operator, and 40% by Japanese partners.The Huckleberry deposit is one of several porphyry systems located in the Tahts a district of the intermontane belt of British Columbia. The principal economic minerals are chalcopyrite and molybdenite. The ore deposit is split into two ore deposits, the Main Zone, consisting of24,241,000 tonnes of mineable reserves grading 0.484 %Cu and 0.013 %Mo, and the East Zone, consisting of 66, 131,500 tonnes of mineable reserves grading 0.523 %Cu and 0.014 %Mo.The milling begins with gyratory crushing. This sizes the product down to 100% passing 6 inches. This product is discharged onto a stockpile by a stationary conveyor. The stockpile was designed for a capacity of 144,000 tonnes with a live capacity of 20,000 tonnes. Two apron feeders draw the ore from the bottom of the stockpile and discharge onto a SAG mill feed conveyor which in turn feeds a grinding circuit consisting of one 32' x 15' SAG mill and two 30' x 16.5' ball mills. The grinding cyclone overflow feeds bulk flotation at 55 -65% minus 200 mesh.The bulk flotation uses conventional rougher and column flotation methods to float a bulk concentrate recovering 92.6% copper on the East Zone and 95% copper on the Main Zone with 75% recovery molybdenum on both zones. The bulk concentrate feeds the bulk thickener which produces a density of 60 %solids. This in turn feeds the molybdenum circuit, which separates the copper and molybdenum, by depressing the copper then using conventional rougher and column flotation produces a copper concentrate of 27 %Cu and 0.02 %Mo, and a molybdenum concentrate containing 1.5 %Cu and 54 %Mo at 67% recovery. These concentrates are separately thickened, filtered, dried, and shipped to the smelter for processing."

Jan 1, 1998

By Francis D. Leathers

Four friction pile load test programs, including 23 tension tests and 9 compression tests, have been performed as part of the Central Artery/Tunnel Project in Boston, Massachusetts. The tests, performed on both precast, prestressed concrete (PPC) piles and steel piles, yielded the following adhesion values: [Soil Type Clay Clay Glacial Till Interbedded Sand and Clay Interbedded Sand and Clay Pile Type PPC Steel Steel Steel PPC Adhesion (psf) 610 to 1,200 230 to 900 220 to 550 680 1,280 ] The adhesion appears to be affected by differences in pile material type, with PPC piles having higher adhesion than steel piles. However, part of the apparent difference for the clay may be due to differences in time between driving and testing. The significant range of test data means that the need for site specific testing to determine or verify design values of adhesion should be considered during design. Creep rates at loads near failure ranged from about 0.02 to 0.18 inches per log cycle of time, with most values in the range of about 0.04 to 0.08 inches per log cycle of time.

Jan 1, 1994

Shaft sinking through an amphibolite-granite rock mass experienced rock bursts on the shaft walls and benches. However, rock burst phenomena was exhibited only in granite, which is e1asto-brittle with ability to store strain energy caused by orogenic and man made stresses. To improve ground conditions, blasting of destressing holes at shaft corners was carried out. By this technique rock burst during shaft sinking was minimized, but not eliminated.

Jan 1, 1982

By G. R. Hopkins

"Refiners are generally considered solely as fuel producers, but actually they also comprise a group one of the largest consumers of fuel. For example, during the year 1929 the refineries-burned 51,544,000 barrels of oil as fuel, or the equivalent of over 5 per cent of the crude oil processed in that year. In addition enough other fuels - coal, natural gas, refinery or still gas, and petroleum coke - were utilized to raise the total fuel consumption based on B. t. u. equivalent to more than 10 per cent of the heat value in the total crude oil run to stills. Refineries utilize fuel chiefly as the heat-forming element distillation, but also under-boilers in the formation of steam used in refining, for the heating of buildings, and in the generation of electricity.In the early days of refining, the problem of fuel utilization and fuel economy was of scant importance to the refiners. Margins of profit were comparatively high and the major portion of the operatros1 attention was directed toward increasing their throughput. However, the keen competition that has prevailed in the refining industry over the past five years or more has taught the refiners that their expenses, one of the largest items of which is the cost of fuel, must be curtailed. Thus in 1925 the refiners used an average of 829,000 B. t. u. to refine a barrel of crude oil, wheras by 1929 this average had been reduced to 639,000 B. t. u., a decline of 23 per cent (see fig. 1)Since 1925 oil has ranked first as a refinery fuel (see fig. 2). The use of natural gas increased during 1926 and 1927 but declined in 1928 and 1929. On the other hand, the use of refinery gas has grown rapidly in importance. Coal has shown a steady decline in use at refineries since 1925. Petroleum coke has never supplied more than a very small fraction of the total fuel requirements, although there has been a steady increase in its use."

Sep 1, 1930

Five watersheds, ranging in size from 29 to 49 acres, were selected in East-Central Ohio to investigate the hydrologic and water quality conditions occurring before, during, and after surface mining. Four of the watersheds will undergo surface mining, involving either the Number 6, 9, 11, or 8 coal seams, and one watershed having an outcrop of the Number 6 coal will be operated as a control watershed. To determine the premining hydrologic and water quality conditions (Phase 1), physical information was obtained for each watershed and instrumentation was installed for obtaining hydrologic data and water quality samples. In addition to topographic maps, soils and vegetation maps were prepared. Stratigraphic columns were developed from geologic cores; networks of observation wells and of soil moisture access tubes were installed; and weather stations were established to measure temperature, humidity, wind and solar radiation, and to collect precipi¬tation samples. Field data were collected over a 10-month period for one watershed and for about 8 months on another watershed before surface mining began. The other watersheds continued in the premining data phase. These data were for precipitation, temperature, relative humidity, wind speed, solar radiation, snow and frozen soil, soil moisture, infiltration (rainfall simulator), spring discharge, runoff from watersheds, and ground-water levels. Also, samples of runoff and ground water were obtained for chemical analyses and additional samples of runoff were acquired for suspended sediment load determinations. Over the study period, precipitation was not very representative of the long term averages, but data are available from the control watershed for a representative 7 year period which should prove invaluable in modeling the hydrology of the project watersheds for the premined conditions. Spring discharges (before mining) from the upper aquifer of two of the watersheds averaged only 508 and 652 gallons daily. The proportion of precipitation occurring as surface runoff from the large watersheds ranged from 8 to 36 percent and the peak discharges from 5.86 to 81.95 cfs. Base-flow and runoff water samples from four of the watersheds (unmined conditions) were collected and analyzed quantitatively for 39 water quality parameters. Monthly precipitation samples sediment fractions of the base flow and runoff were analyzed for 22 and 17 parameters, respectively. The average concentrations of the water quality parameters for the watersheds (unmined), except for suspended solids, were as low or lower than the EPA recommended maximum allowable concentrations in drinking water. Several parameters in precipitation equaled or exceeded the levels of corresponding parameters in runoff water. The average concentration of suspended solids in storm runoff ranged from 118 to 1110 mg/l, and for base flow the range was from 44 to 247 mg/l; the higher concentrations were associated with the watersheds exhibiting the greatest disturbance. Iron ranged from an average of less than 3 µg/1 to 228 µg/1. Manganese ranged from less than 41 µg/1 to 334 µg/I. Ground-water systems (unmined condition) are described for the five project watersheds. Each watershed was found to have two major perched aquifers within the top 250 feet. Generally, these perched aquifers involve local flow systems, though flow across watershed boundaries in some areas. Recharge is mostly from local precipitation and discharge is mostly as spring flow, base flow, evapotranspiration, and leakage through the underlying confining bed. Shallow ground water was commonly of the calcium bicarbonate type. Although dissolved-solids concentration was generally 200 to 600 mg/1, brackish or salty ground water was present in several areas. A composite hydrologic model for the project watersheds will consist of separate ground-water, unsaturated flow, overland flow, and evapotranspiration component models. The ground-water model (a modified quasi-three--dimensional U. S. Geological Survey model) operates on a grid network and the newly formulated unsaturated flow model was divided into overlying prismatic elements for numerical solutions. The unsaturated flow model includes provision for incorporating the evapotranspiration model. An available distributed kinemetic overland flow model was utilized. An analyses was made of one of the costs of reclaiming coal mines; the legal and administrative costs of complying with legislation. Such costs borne by coal companies under the Ohio law are thought to be a surrogate of these costs under the new Federal law. Regression analyses were used to relate the legal and administrative costs to acres in the permit and the tonnage of coal produced.

Jan 1, 1978

By W. D. Monaghan

Flame cutting and welding is one of the major causes of fires and explosions in underground U.S. coal mines. On May 20, 2006, a flame cutting operation at the Darby Mine No. 1 led to an explosion that resulted in five fatalities. On January 22, 2003, another explosion caused by flame cutting at the McElroy Mine resulted in three fatalities and three serious injuries. On June 22, 1999, a flame cutting and welding operation at the Loveridge Mine caused a fire that resulted in sealing of the mine for an extended period. Finally, on March 19, 1992, an explosion at the Blacksville No. 1 Mine that was caused by welding resulted in four fatalities. The National Institute for Occupational Safety and Health (NIOSH) conducted a study on fires and explosions in underground U.S. coal mines that were caused by flame cutting and welding operations to determine the root causes of these types of fires and explosions. The methodology included interviewing mining personnel who perform flame cutting or welding operations in underground U.S. coal mines and visiting mines to observe these operations, and MSHA reports of investigations and accidents statistics were analyzed. The results were used to identify and compare differences between flame cutting and welding practices and techniques in small and large mines, Eastern and Western mines, low seam and high seam mines, room and pillar and longwall mines, and between experienced miners and new miners. As a result of this study, best practices and interventions were developed to reduce the number of fires and explosions caused by flame cutting and welding. This paper provides a summary of this work. Disclaimer: The finding and conclusion in this report are those of the authors and do not necessarily represent the views of the National Institute for Occupation Safety and Health.

Jan 1, 2009

By Eleanor Sonley

In October 2000, temporary seismographic stations were installed above the Trail Mountain Mine, Emery County, Utah, USA to supplement a regional network. Over seven months, 1826 mining-induced seismic events were recorded with a magnitude range of M 0 to M 2.2 (Arabasz et al., 2002, 2005). Routine event locations for this dataset cluster around the longwall mining activity during the recording interval. However, applying a double difference relocation technique to the routine locations significantly improves the correlation between mining activity and event location. The resulting relocations depend greatly on station distribution and the number of arrival time picks. We use this dense temporary array to assess station distribution requirements for applying the double difference technique at other mines by systematically removing data from available stations, thus simulating sparser networks. The insight gained using the Trail Mountain dataset is used to develop a metric for assessing the usefulness of above-mine networks and to develop a tool for expanding seismic network coverage. To verify the appropriateness of these tools for wider application, we examine data from two additional mines located in the western United States. Improving the relative locations of mining-induced seismic events highlights the association of these events with active portions of the mine. Clusters, regions of quiescence and relative timing of events can be used by mine operators to assess hazard, optimize production, and evaluate performance of mine plans. The lessons learned from this study may be used in future network design to improve accuracy of event locations at other mines.

Jan 1, 2009

By S Thomson, M Lovitt, T Purvis, C te Kloot, T Nguyen, Z Liu

Ernest Henry Mine (EHM), located in north-western Queensland, is an underground copper/gold sublevel caving operation currently producing over 6 Mt/a of ore. The mass mining method of sublevel caving (SLC) used at EHM involves systematically blasting a series of uphole rings in parallel ore drives and drawing the broken ore using a top-down approach. Each ore drive contains multiple precharged rings to prevent having to charge in close proximity to the brow of the previously fired ring; however, the current initiation system still involves ‘hooking up’ all detonator lead cords of the first precharged ring. The unstable nature of a rill and working close to a brow presents hazards to personnel completing the hook-up process. There is also a business risk of ore loss when damaged ground support completely prevents access to the next precharged ring. If the hook-up step could be eliminated, the risk to personnel and the business would be significantly reduced.EHM engaged Orica Mining Services in 2016 to trial their WebGenTM 100 – wireless electronic blasting system (WEBSTM) for the first time in an underground mining environment. Having successfully signal-tested the equipment, two test holes were initiated, followed by a single production ring and then multiple production rings. In total, 30 production rings were successfully fired underground at EHM in 2017. To do this, WEBSTM was successfully integrated into EHM’s current charging and firing process, resulting in no business delays or loss to production. None of the 30 rings required hooking up and all were initiated from the surface with no personnel underground. A high level of ownership from all personnel involved ensured the trial was a success, and provided scope for wireless detonators to be used successfully on a large scale – both at EHM and other sublevel caving operations. WEBSTM is now a proven example of technology that can be used to improve safety in the mining industry, through eliminating the need for personnel to work in hazardous areas.CITATION:te Kloot, C, Liu, Z, Purvis, T, Thomson, S, Lovitt, M and Nguyen, T, 2017. Improving safety through technology – Orica wireless electronic blasting system trials at Ernest Henry Mine, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 39–50 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 16, 2017

Kinetic Study on the Leaching of Vanadium from LD Converter Slag Using Sulfuric Acid In order to reduce the environmental impact due to land disposal of LD (Linz-Donawitz) converter slag from steelmaking plant and to valorise this waste material, an acid leaching study was carried out to assess the effect of several parameters on the kinetics of vanadium dissolution. The characterisation and morphology of the feed investigated using XRD, XRF. The factors influencing leaching vanadium such as particle size, temperature, slurry density and the concentration of sulfuric acid were investigated and optimised. The maximum vanadium recovery about 95 per cent achieved at 70¦C, S/L: 1/15, acid concentration: 3M and for 150 minutes of leaching time. The dissolution of vanadium in sulfuric acid shows that there are two stages in the kinetics of leaching. At the first 15 minutes, there is a sharp increase in the amount of vanadium extracted and at longer leaching time, the extraction increase slightly. In the present work, the shrinking core model (SCM) was used to describe the acid leaching of vanadium compounds. A summary of the controlling regimes in the SCM are examined below for short time (less than 12 min.) and long time (more than 30 min.) stages. So, it was found that the kinetics of leaching in both stages is controlled by solid product diffusion regime.

Sep 13, 2010

By N. D. L. Burger

Underwind and overwind protection system concepts for mine hoist shafts were developed in conjunction with and for the Safety in Mines Research Advisory Committee (SIMRAC). End-of-wind operation of mine hoists is the most hazardous aspect of mine hoisting, carrying the highest risk of loss of life, injuries to workers and loss of production. Systems operating mechanically and with local actuation were developed. This was done in order to remove the increased risk associated with dependency on remote or external energy and information supply associated with most existing protection systems. The proposed underwind protection system concept absorbs the energy of motion of a conveyance overrunning the design lower limit of travel by drawing a metal strip through a set of rollers. This action causes dynamic cyclic plastic bending of the strip material that converts the kinetic and potential energy of the conveyance into strain energy of the metal. In the concept design a pair of steel wire rope slings attached to the strips catch the overrunning conveyance and transfer the retardation force. The proposed overwind protection system concept absorbs the energy of motion of a conveyance overrunning the design upper limit of travel by early detaching of the conveyance from the hoist rope. This detaching is carried out via an additional detaching hook activation mechanism fitted at a sufficient height below the spectacle plate to allow the conveyance to retard to standstill under gravity before it would crash into the spectacle plate. The conveyance so brought to rest is prevented from falling by means of jack catches on the conveyance interacting with a rack (toothed profile) fitted on the guide rails in the retardation zone. Scale models (1:10 scale) of both protection systems were designed to conform to established retardation standards. The retardation standards were limited by the requirement of passenger safety in retarding cages. The models were then built and tested in a1:10 scale shaft model. Retardation performance close to the required levels was achieved. The retardation distance required should allow such systems to be retrofitted in most existing shafts. Given the simplicity and robustness of the designs, further development was recommended because of their enhanced self-sufficiency and reduced risk of malfunction. Keywords: underwind, overwind, deceleration, mineshafts, hoisting.

Jan 1, 2004