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By R Sharpe, M D. Hannington, K Perrin, K Simpson, S Petersen, H Gibson, N Augustin, K Lackschewitz, T Monecke

Semi-massive to massive sulfides were drilled in a shallow-water (630 m) hydrothermal system in an island arc volcanic setting at the Palinuro volcanic complex in the Tyrrhenian Sea,Italy. Overall 12.7 m of sulfide minerlisation were drilled in a sediment-filled depression in a water depth of 630 - 650 m using the lander-type Rockdrill 1 drill rig of the British Geological Survey. The longest drill core recovered consists of 4.9 m of continues semi-massive to massive sulfides + sulfates with abundant late native sulfur overprint. Metal enrichment at the top of the deposit is evident in some cores with polymetallic (Zn, Pb, Sb, As, Ag) sulfides overlying more massive and dense pyritic ore. The massive sulfide mineralisation at the Palinuro volcanic complex contains a number of unusual minerals, including enargitefamatinite, tennantite-tetrahedrite, stibnite, bismuthinite and Ag-sulfosalts, that are not commonly encountered in mid-ocean ridge massive sulfides. In analogy to epithermal deposits forming on land, the occurrence of these minerals and the abundance of native sulfur suggest an intermediate to high sulfidation state for the hydrothermal fluids during deposition, implying that the mineralising fluids were acidic and oxidised rather than near-neutral and reduced as those forming typical base metal-rich massive sulfides along mid-ocean ridges. Oxidised conditions during sulfide deposition are likely related to the presence of magmatic volatiles in the mineralising fluids that were derived from a degassing magma chamber below Palinuro volcanic complex. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By Poole GR

An unusually high flow of water from an oxtr.iction panel in Wongawilli Colliery pr,;npted the investigation of the possible sources of the water. It is more probable that water is derived f r,m sources adjacent to or within the Woo ' iwi 1 I i Coal rat her than by direct t r.onnuni n i"n from any surface source.

Jan 1, 1986

By Dunbavan M, Mallett C. W

The sequence of development and back analysis of the failure of a spoil pile in the 1B pit of the Moura Mine are given. The mode of failure was a two-wedge mechanism with a basal slip plane passing through a clay band about 2 m below the pit floor. Remedial actions for improving spoil pile stability are discussed.

Jan 1, 1981

By A Holmgren, W Forsling

In the present study infrared attenuated total reflection (IR-ATR) was used to monitor the rate of in situ adsorption of heptyl xanthate on a layer of zinc sulfide synthesised on a germanium ATR crystal. The zinc sulfide surface was characterised using x-ray photoelectron spectroscopy (XPS). The absorbance of heptyl xanthate measured increased with increasing bulk concentration of the adsorbate up to an equilibrium plateau value corresponding to a fractional coverage of the surface sites of the substrate. For the adsorption from solutions of concentrations between 10-3 and 10-5 M, the rate of adsorption increased with increasing bulk concentration. At higher concentrations the measured absorbance should be corrected for the absorbance due to the concentration of xanthate in bulk solution. The present work provides a discussion of this correction. Assuming the adsorption stage to be rate-controlling and proportional both to the concentration of the adsorbing species in solution and unreacted surface area, a rate law governing this type of transient adsorption kinetics was derived. By means of absorbance data for the concentrations 0.1 and 0.01 mM, the rate constant of adsorption was calculated. From the initial part of a Langmuir adsorption isotherm, the absorbance at monolayer coverage and the equilibrium constant of adsorption could be estimated.

Jan 1, 2005

Most coal mines in Chongqing Area are highly gassy mines and liable to outburst of coal cnd methane. In order to harness mine environment and raise the Prodction capacity, methane predrainage is generally taken as a main preventive measure against outburst by all mines. In the recent ten years, the drainage volume per year has increased from 38.25 million m3 to mote than 100 million m3 with a volume over 12 m3 per tome of coal. In some mines, the draining rate is approaching to 50 percent, the rate in mining districts is up to 60 to 70 percent. The predrainage volume accounts for one fourth of total drainage volume. All these have effectively prevented the occurrence of coal and methane outbursts.This is the current situation of methane predrainage in coal mines in Chongqing area, the methods of predrainage to prevent outbursts and related parametres comprise:1. methods, technologies, supervision and effect of predrainage in different geological and mining conditions,2. the change of methane parametres pre-and post-drainage and the analysis of the effect of methane predrainage to prevent outbursts, and3. ways to improve the effect of methane predrainage.Thus the safety, reliability and production of Changqing Coal Mines will be enhanced.

Jan 1, 1991

By Fukuichi K, Matsukawa H, Nakagawa M

The Myojin Tunnel, located on the Shikoku Expressway being constructed by the Japan Highway Public Corporation, is 3,770 m long with a cross sectional area of 80 m2. Kajima Corporation, the contractor for the northern construction section of this tunnel, adopted here its newly developed "mini-bench" method of excavation. The objectives of this method were 1. to be able to perform all tunnelling operations using the whole cross sectional area of the tunnel under any ground conditions, and 2. to be able to vary the length of the mini-bench within the range of 0 to 7 m according to ground conditions. In order to achieve these objectives, original, sophisticated equipment was developed and employed in the work. One of the newly developed machines is a special jumbo, a hydraulic crawler jumbo with five rotary type booms. This machine is capable of accommodating variation in the length of the mini-bench accompanying changes in ground conditions and is equipped with a special mechanism by which the positions of the booms can be controlled by means of hydraulic cylinders. Also developed was a shotcrete robot for use with the mini-bench method which has the control board near the shotcrete nozzle for close control of application.

Jan 1, 1987

By P S. Jaitawat, M Sarim, M P. Roy, P K. Singh, A Joshi

Mining is being carried out using the sublevel open stoping method at the Kayad underground mine. The mine will ultimately reach a depth of 390 m. Kayad village is only 100 m (horizontally) from the underground mining areas. As a result, blast-induced ground vibration is one of the major issues for the residents of the village. This paper deals with the process involved in optimising blast design parameters at the development face blast, slot raise blast and ring hole blast to control blast-induced vibration within safe limits while ensuring improved production, productivity and safety. In total, 1041 development face blasts, 110 slot raise blasts and 52 ring hole blasts were conducted, and 2612 vibration data were monitored in the village at different locations. Initially, electronic delay detonators were used, with an emphasis on controlling vibration in the surface structures. Subsequently, shock tube (Nonel) initiations were used as the blasting faces reached lower depths. A comparative study was also performed for both the initiation systems. The orebody width generally varied from 3.1 to 7 m, and ring hole designs were optimised accordingly. The number of holes in a ring was up to seven and the depth was up to 23 m, depending on the size/width of the orebody. The ring blasts were performed with 64–340 kg of emulsion cartridge explosive and detonated with explosive weight per delay of 10.92–17.16 kg. Extensive experiments were performed and blast-induced vibrations were controlled in the village. Deck blasting (ie up to four explosives decks) was used in a blasthole and the explosive charges were separated by delay detonators, which helped to control the vibration in the village. The actual detonation time of each deck was recorded by monitoring the in-hole velocity of detonation of the explosives, and the charging patterns were optimised accordingly to get the desired blast results.CITATION:Roy, M P, Singh, P K, Sarim, M, Jaitawat, P S and Joshi, A, 2015. Blast design and vibration control at Kayad lead-zinc underground mine, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 533–544 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By H Hoogvliet

The Goondicum Crater industrial minerals deposit is located about 30 km east of Monto, Central Queensland, Australia. Early exploration programs in the 1990s found alluvial ilmenite along the banks of the Burnett River. This mineralisation was traced upstream to the Goondicum Crater. The Goondicum Crater is a topographic feature centred on a roughly circular 6 km diameter layered gabbro complex. Significant ilmenite, apatite, titano-magnetite and feldspar mineralisation are present in a weathered horizon up to 25 m thick. After several drilling campaigns in the Crater and subsequent feasibility studies, the project was developed by Monto Minerals Ltd, an ASX listed company, in 2007. However, the project never achieved production performance expectations and ran out of commissioning funds. The company subsequently went into voluntary administration in August 2008.In 2009, the RMM Capital group, of Brisbane, took control of Monto Resources Pty Ltd, and renamed the company Belridge Enterprises Pty Ltd (Belridge). Belridge commissioned a redevelopment study aimed at recommencing mining and processing operations. As part of the redevelopment study the authors planned a drilling campaign, managed a quality control program and developed a block model. This paper describes the drilling program, analytical procedures, quality control procedures and the preparation of the block model.The block model for the resource estimate was created using 1 m high blocks, wireframes defining the main material types, and 2.5 m thick horizontal search ellipsoids with a radial pattern. The deposit was ‘flattened’ prior to modelling. To assist with the economic evaluation of the project, each block in the model was assigned a net dollar value based on economic parameters provided by Belridge, including anticipated revenues and costs. This modelling technique allowed for efficient investigation of various scenarios and was used extensively in the development and optimisation of the project’s financial model. At a zero net value per block, comparable to a break-even cut-off grade of four per cent ilmenite equivalent, the Indicated plus Measured resource is estimated at 51 Mt, averaging 4.1 per cent ilmenite, 1.9 per cent apatite and 2.5 per cent titano-magnetite.

Oct 5, 2011

The Teutonic Bore deposit lies in a major greenstone belt within the Archaean Yilgarn craton of Western Australia. It occurs towards the base of a suite of tholeiitic basalts, some 100 m above a felsic volcanic pile of rhyolitic composition. Sediments comprise a minor, albeit important, percentage of the mine stratigraphy.The deposit consists of a single, steeply dipping lens of massive sulphides totalling 1.4 million tonnes grading 4.16 per cent Cu, 16.4 per cent Zn, 1.22 per cent Pb and 203 g/t Ag. In addition, cross cutting "stringer" mineralisation in the footwall contributes a further 0.75 million tonnes at 2.38 per cent Cu, 1.92 per cent Zn and 52 g/t Ag to the mining reserves. Principal primary sulphides in massive and disseminated mineralisation are sphalerite, pyrite, chalcopyrite and galena, while minor amounts of arsenopyrite and argentiferous tetrahedrite are also present. Weathering has affected the primary mineralogy down to at least 90 m, resulting in a strongly leached oxide profile to about 85 m, overlying a thin, erratic zone of chalcocite supergene enrichment.Lateral metal zoning is marked both in massive sulphide and stringer mineralisation, with Cu/Zn ratios decreasing away from the centre of the orebody. Alteration is well developed in association with footwall stringer mineralisation, changing from a central Fe-Mg carbonate, chlorite and silica alteration zone below the massive sulphides to an outer halo of sericitisation. The spatial distribution of these minerals is patchy. Geochemically, the footwall alteration zone is characterised by central Cao, Na20 and Sr depletion (with less regular enrichment of SiO2 , Fe and MgO), surrounded by a peripheral zone of K20 enrichment.Evidence supports a classic volcanic-exhalative origin, involving submarine discharge on the flanks of a rhyolitic dome and accumulation of the massive sulphides in a local palaeotopographic depression. The sulphides overlie their feeder zone as shown by alteration and stringer mineralisation. Tholeiitic wallrocks are thought to be genetically unrelated to the mineralising solutions, although their presence as a cap over the felsic volcanics may have permitted the build up of a circulating metalliferous brine system prior to sulphide deposition.

Jan 1, 1984

By M C. Ackland

The Grande Côte Mineral Sands Deposit (GCMSP) is located on a coastal mobile dune system in Senegal on the west coast of Africa. Senegal is a stable, democratic republic, having gained its independence from France in 1960. It has a population of about 12.5 million people. The dunes begin about 25 km north-east of Dakar, the capital city of Senegal, and extend northwards for more than 140 km. The mineralised dune system averages about 2 km wide with some areas extending to up to 4.5 km wide.Mineral Deposits limited (MDL), a Melbourne-based company with over 60 years operational experience in mineral sand mining in Australia, holds, through its Senegalese subsidiary MDL Senegal SARL (MDLS), a 25-year Mining Concession covering the deposits. The main heavy mineral (HM) areas identified to date are, from south to north, Mboro, Fass Boye, Diogo and Lompoul. Within the Mining Concession there is potential to identify additional resources beyond the limits of present drilling. The dunes are mobile or semi fixed, pale yellow in colour and overlie older white marine sands. The dunes range between 5 m and 30 m in height and the mineralised zones, which are essentially flat-lying, average around 15 m in thickness. The mineralisation consists, principally of ilmenite with accessory zircon, rutile and leucoxene. Zircon and ilmenite are the main commodities of interest.Independent consultants AMC estimated mineral resources for the four main deposits in 2006 of 1330 Mt of sand averaging 2.0 per cent HM at a cut-off grade of 1.5 per cent HM. At a cut-off grade of one per cent the resources were 3200 Mt averaging 1.5 per cent HM. All of the estimate was classified as Inferred Resources under the 2004 Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves (JORC, 2004).Subsequent completion of infill drilling, sampling and assaying of the initial mining area, centred at Diogo, was completed in 2008, enabling a final dredge path to be designed and ore reserves estimated, based on a JORC compliant Measured and Indicated Mineral Resource. Infill drilling is continuing in the Lompoul and the results of this campaign will be included in the JORC compliant data base by AMC in April 2010 to enable the 15 per cent year dredge path and mine plan to be finalised.Development studies have shown that the project can be optimised as a dredging operation, with mineralised sand being treated in a conventional floating spiral preconcentrator, wet concentrator and dry mineral separation plant. Engineering work now substantially completed by Ausenco Limited (Ausenco) has developed the combination of some existing plant from MDL’s past operations, and new plant for the complete process from dredging to final product separation and grading. Testwork on a series of bulk samples has determined that the project can yield a high quality zircon product and an ilmenite product along with minor rutile and leucoxene.Work to date indicates a likely output of around 80 000 t of zircon in various quality streams (Primary, Secondary, Standard and Foundry Grades), around 550 000 t of ilmenite plus around 15000 t of rutile and around 12 000 t of leucoxene/secondary ilmenite. Planned zircon output would represent around 7per cent of the total world production, which would make the GCZ&IP a significant producer on a world scale.The studies have indicated capital costs to be in the range of US$400 M total operating cost (after completion of ramp-up) of around US$50 M - $55 M/a. Revenue is expected to be in the range $127 M - $157 M/a.

Oct 5, 2011

By Ueki S

Underground coal gasification (UCG) is a promising process for obtai- ning energy from coal, which is either unmineable with current mining techno- logy or uneconomic to recover, using conventional methods, in the present energy market situation. There are several coal fields, which could become new coal reserves, if the UCG process were successfully applied. In the 1960's Japanese researchers made a number of basic studies and field tests into UCG, although not all of them produced favourable results. This paper reviews Japanese research on UCG. This includs basic studies into coal gasification using a small, tube reactor fed by air and oxygen, and also large scale model stu- dies(reactor size of 8.Om x 2.4m x 1.5 m). Field test results from Akabira colliery in Hokkaido and a feasibility study of power generation using UCG from a coal field in Hokkaido are also Presented.

Jan 1, 1987

By Fukui M

An increase in pig iron output by 50 per cent was aimed at after re-lining No.2 blast furnace of NAKAYAMA STEEL WORKS. This had to be achieved by increasing both the furnace volume (from 673 to 757 m3) and its productivity, Without equipment for high top pressure operation, high productivity was mainly attained by the following three measures; (1) improving the gas flow characteristics by maintaining a high ratio of agglomerated ores, increasing the strength of sinter and coke, and optimizing the charge distribution, (2) decreasing the fuel ratio by low Si operation and the use of hot blast, and (3) oxygen enrichment. In addition, the tapping controls for pig iron and slag were intensified. As a result of the above measures an excellent hot metal output of 2.7 tonne/m3/day was attained in April 1979 (monthly average).

Jan 1, 1980

By Mowat B. A

I lamata is within the Morobe Goldfield (7¦ 25' S, 146¦ 39' E) on the Wau (SB 55-14),1:250,000 scale map sheet (Fig 1). The deposit is 13 km SW of Wau, within the headwaters of the Upper Watut River, at an average elevation of 2000 m. The steep mountainous terrain is covered by dense tropical rainforest and access is by helicopter or by foot track. The deposit has been described previously by Wells and Young (1991) and 1)enwer. Leach and Mowat (1995). Data collected to late 1996 have been used to estimate an Inferred Resource of 9.2 Mt at 3.1 g/t gold using a cutoff of 0.5 g/t. This contains a high grade portion for which an Inferred Resource of 1.2 Mt at 9.7 grt gold has been estimated using a 5.0 g/t cutoff.

Jan 1, 1997

By Bratt G. C, Horsham T. R, Langlois D. R

Mercury is precipitated from weakly acidic sulphuric acid plant waste liquor at the Risdon Works of the Electrolytic Zinc Company of Australasia Limited. Factors determining the choice of process and reagent, as well as details of process development including scale- up, and the time scale of implementation are outlined. The plant is designed for a maximum flow of 120 m3/h.of liquor containing mercury and saturated with sulphur dioxide. Mercury is precipitated by fluidised beds of aluminium pellets contained in lead columns. Liquor is then clarified with plate and frame presses. Experimental work on the mercury precip- itation reaction was conducted in three phases involving a scale-up of 100 times in equipment size. Extensive pilot filtration work was carried out on a 1 m2 press. Scale-up of 120 times was required for specification of plant filters. The basis of scale-up is outlined.

Jan 1, 1977

By G R. Hancock, M J. Saynor, J B. C Lowry, K G. Evans, M Narayan

Tools such as landform evolution models (LEMs) provide a means of predicting how a landform may evolve over extended periods of time. Here, we utilise the CAESAR-Lisflood LEM to assess a proposed rehabilitated landform of the Ranger Uranium Mine in the Northern Territory of Australia (Figure 1). The environmental requirements for the closure of the mine state that the final landform must be able to physically isolate tailings material from the environment for a period of at least 10,000 years (Australian Government, 1999). However, uncertainty exists as to the reliability of model predictions representing thousands of years.In this study, we use model predictions from an undisturbed pre-mine landform to calibrate the parameters used in model simulations of a post-mining rehabilitated landform over a simulated period of 1,000 years. Specifically, we vary environmental conditions such as rainfall, presence/absence of vegetation and surface particle size characteristics in a series of model simulations to determine their impact on the respective landforms, to revise model input parameters as required. A simulated period of 1,000 years was used as it represents a period in which long term landform evolution patterns and trends may be established and identified. CITATION:Lowry, J B C, Narayan, M, Evans, K G, Saynor, M J and Hancock, G R, 2018. Using an undisturbed landform to calibrate long-term predictions of the evolution of a rehabilitated landform, in Proceedings Life-of-Mine 2018, pp 133–137 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 25, 2018

By Smith N. S, Rholl S. A, Otterness R. E

The Bureau of Mines is conducting research to develop blast designs for the control of fragmentation and rock displacement and the reduction of overall mining costs. This Bureau experiment, a program of 25 reduced-scale blasts conducted at the University of Missouri-Rolla's experimental quarry, determined the effects on fragmentation of changes in burden, spacing, explosive diameter, and stemming-explosive length. Combined with previous work at this experimental quarry examining air gap, delay, and decoupling effects, this experiment provided a data base of over 50 shots. The fragmented rock, 1000 to 3000 kg from each 3- to 4-hole blast in the 1 m bench of massive dolomite, was completely screened to assess fragmentation. The explosive used for all the tests was extra dynamite initiated by seismic caps fired from a sequential blasting machine with delays of 5.2- to 13-ms/m of burden. Empirical equations were developed which predicted the effect of shot design parameters on the fragmentation. These formulas predicted the average fragment size of each of the 25-shots to within 13 percent. For spacing to burden ratios of 1, 1.4 and 2, there were optimum burden to explosive diameter ratios (B/de) of 30, 26 and 20, respectively. Decreasing the B/d. ratio beyond these levels resulted in little or no reduction in the average size. For B/de ratios above these levels the average size increased but, at least a 15 percent reduction in the average size results when the shot design parameters are optimized.

Jan 1, 1990

By Ryan A, Dennis R, Dekker D, Ellis W, Sowerby B

An automated facility for non-destructive elemental analysis of mine samples has been developed and tested at Mount Isa in association with the CSIRO. The analysis is based on the prompt and delayed measurement of gamma radiation emitted by the sample during and after separate fast and thermal neutron irradiation. No sub-sampling or sample preparation is required and the analyser is operated by one person that would otherwise be splitting drill core in preparation for conventional analysis. The analyser has been calibrated to measure the following (precision in parenthesis): Ag (14 ppm), Pb (0.5 wt%), Zn (0.5 wt%), Cu (0.04 wt%), Fe (0.7 wt%), S (1.1 wt%), SiO2 (0.8 wt%), A1203 (0.07 wt%), Mn (0.05 wt%), specific gravity (0.02 g cmt). Crushed or whole drill core samples that range in size between 0.5 m and 1.8 m of equivalent BQ size drill core can be analysed at a rate of six samples per hour. If required, drill core samples can be re-packed in their original configuration for storage following analysis. Standard samples are measured every day to ensure operational stability. The analyser is controlled by a personal computer which prints the results and stores them along with the raw data on the computer hard disk. This paper reports the performance of the analyser during 1992 for routine analysis of drill core from the lead ore bodies of both Mount Isa and Hilton mines.

Jan 1, 1993

By M Nagarajan, S Esen

This paper presents three case studies from three surface coalmines. The first case study is from a coalmine where a dragline operates in 60 m wide and 1 km long strips. Typical overburden height ranges from 30 m to 50 m. The base case blast was modelled using an advanced blasting model (Distinct Motion Code – DMC) to establish the baseline using the blasting parameters in use before optimisation. DMC was used to model the alternative blasting scenarios to obtain the following key objectives:the post-blast muck pile profile was required to have an optimum height of ~28 moptimise cast percentageachieve a well-fragmented muck pile for the draglinethe muck pile profile should be appropriate for the dragline entry into the strip.Achieving these objectives would optimise the dragline performance through reducing rehandle and improving advance along the strip. DMC modelling resulted in the following outcomes:the required muck pile profile could be achievedthe cast percentage increased from 21.1 to 25.1 – a four per cent increaserehandle volumes were reduced.Establishing a good quality assurance / quality control (QA/QC) process was essential to delivering a blast that could achieve these outcomes. The blast was fired successfully and the dragline completed the strip approximately two weeks ahead of the time taken for the previous strip.The second case study is from a coalmine where D11 dozers are used to move the blasted material to the spoil. An optimum profile was required for the dozers for this low-cost mining operation. To achieve this, an electronic blasting system (uni tronicTM 600) was introduced to optimise the muck pile profiles being delivered. The new muck pile profiles showed improved cast percentage from 10.6 per cent to 17.8 per cent. The changes in the muck pile centre of mass displacement improved dozer efficiency as a result of pushing material shorter distances to its final position. Dozer productivity (bcm/h) statistically did not change but coal was being exposed at a faster rate.The third case study summarises the results from a site where dozers are used to minimise the total mining cost. Traditionally, excavators had been used on-site with conventional blast designs. After firing the first cast shot, the blasted material was found reasonably well-fragmented. Although dozing productivity hasn’t been quantified on-site, operators were providing positive feedback (easier push) due to the reasonably good muck pile profile suiting to the dozers. This site also realised significant savings with the dozing as compared to their truck/shovel operation.CITATION:Esen, S and Nagarajan, M, 2015. Muck pile shaping for draglines and dozers at surface coalmines, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 409–416 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By K Hortin

Kidd Creek mine in north western Ontario is developing phase 1 of the deep mine project for mining between 2100 m and 3100 m below surface. Heat problems were experienced during the 2003 summer in working places about 2300 m below surface. At the time the mine had a æcold stopeÆ for cooling the air supplied to the mine below 1500 m from surface. The æcold stopeÆ is where during winter, ice is created in both the open pit and its connecting airways to the mine and used to cool air in summer. The operation and limitations of the existing cold stope were evaluated and led to the requirement for additional mechanical cooling to be available for the summer of 2004. The first stage of a surface refrigeration plant providing 7.5 MWR of cooling was approved, the compressor set was ordered in early January 2004 and the plant commissioned six months later in early July 2004. Based on the current modifications to the mining plan, further plant expansions are required in 2006 and 2009 leading to a total of 19 MWR of mechanical refrigeration and 8.5 MWR equivalent from the extended cold stope. The refrigeration system is described with emphasis on the specific adaptations necessary to suit Canadian conditions. Additional investigations were; using the refrigeration compressors in winter for heating and, extending the performance of either cold stopes or alternative ice storage systems. The winter use of refrigeration involve removing heat from the mine exhaust air to provide a cooling load and to use the hot water in the condensing circuit to heat the intake air and reduce reliance on mine air heating with natural gas. This paper follows on from an earlier one presented at the Seventh International Mine Ventilation Congress where the background and planning of the mine ventilation and cooling requirements were described for extending the mine at depth to 3100 m.

Jan 1, 2005

By J Player, P Mikula, G Sweby

A numerical modelling back-analysis exercise was carried out at a WA gold mine at which early signs of seismicity (rock noise, scatts from active faces) were becoming apparent as depth of mining increased towards the 500 m mark. Some strain bursting had occurred and it was necessary to discover what the implications of this were for greater depth. The aim of the modelling was to: 1. Establish a correlation between observed ground conditions and model predicted damage, in overstressed areas of the mine. 2. Use this correlation to forecast rock mass damage and ground control requirements for proposed future stoping panels at greater depth. Good correlation has been achieved between observed rock mass damage, and a 3D Finite-Element RS3 model, run in elastic mode. Two candidate stress field orientations were modelled. The model parameter best fitting the observations was the Strength Factor (SF), the ratio of strength/stress. The good match (correlation coefficient of 0.7) between observed and modelled suggests that the elastic approach is justified for the depth range and rock mass properties analysed. The correlation also enabled a link between observed damage and installed ground support performance. Using the linear best-fit between observed and modelled, a forecasting exercise was carried out using the same RS3 model, modified to include future stoping. The resulting damage and ground support categories were obtained from the modelled SF for each proposed stoping level. Also, a series of generic geometries was assessed to investigate the impact of varying stope face lead/lag distance, sill pillar height and stope pillar size, on ground conditions. The key finding of the generic geometries were: • Ground support schemes have been defined for the expected conditions. • All ore drives are expected to have significant damage in the backs and rehabilitation. • Lead/lag distance not to exceed 10 m. • Sill pillar/level intervals of less than 20 m require an increase in ground support scheme. • Pillar sizes for maintaining footwall and hanging wall stability should be at least 10 m × 5 m.

Nov 29, 2022