Search Documents

By M Dixon, T OÆShannessy

Melbourne Water (MW) is responsible for a section of existing gravity trunk sewer servicing the Melbourne central business district, Docklands, Port Melbourne and other adjacent areas, known as the Melbourne Main Sewer (MMS). MW has determined that the 2.2 km long 110 year old sewer will not adequately service these areas into the future or provide the level of reliability needed. The Melbourne Main Sewer Replacement project (MMSR) involves the replacement of the MMS with a new gravity sewer main and six key manhole shafts, known as the Melbourne Main Reliever (MMR).The majority of the MMR sewer tunnel is situated south of the Yarra River and will be constructed using an earth pressure balance (EPB) tunnel boring machine (TBM) and will be lined with precast concrete segments. The tunnel will be 2.4 m internal diameter and will have a 1.8 m diameter glass fibre reinforced plastic (GRP) sewer pipe positioned within it. This project represents the first large-scale, modern, mechanised tunnel project in MelbourneÆs Yarra Delta Quaternary sediments. There is little precedence known to the authors of similar scale tunnelling projects in these particular geological units. During the recent detailed investigation and design stages of the project, significant effort has been spent on characterising the geotechnical conditions along the alignment and predicting the effects of the new tunnel on the existing environment. Some of the key design issues that have been identified and managed through the design process include short-term surface settlement due to tunnel excavation and the predicted effects on surface infrastructure, potential spalling of the steel fibre reinforced concrete segments under full TBM ram loads and tunnel lining design in geological units with highly variable stiffness.Construction of the new MMR sewer tunnel is expected to commence in mid-2008 with a capital cost of approximately $160 million. Sewer system commissioning is scheduled for completion by early 2012.

Jan 1, 2008

Teleoperation is one method by which an operator can remotely control and monitor a vehicle. The operator monitors the status and progress of the remote vehicle by watching the video transmitted from one or more cameras which are mounted on the vehicle. In the case of an LHD, it will be fitted with three cameras. For a twin-boom Jumbo, four cameras are mounted on the drill, three of which will have remote pan, tilt, 15x zoom, focus and lens cleaning controls. The advantage of teleoperation is that an operator can stand in the drive and send the remote vehicle around a corner, completely out of sight and still clearly see what is happening. This means that an LHD returns with a full bucket and a Jumbo can be operated from a safe distance. Standing in the drive teleoperating vehicles, exposes the operator to passing traffic including the remote vehicle. There are two solutions to this problem. One is to equip the operator with a Nautilus æBuddyÆ jacket which will help to protect them from all vehicles. The other is to move the operator into a safe room or area, away from the fumes, the noise and the traffic. In this comfortable, less stressful environment, the operator will be more productive and less likely to accidentally damage the vehicle. The operatorÆs æsafeÆ room does not have to be a formal control room, it can be any area which has AC or DC power available and is large enough to contain a chair and a 1.5 m x 0.75 m table. The æsafeÆ room can be on the surface or underground, it does not matter to the operation of the Nautilus Equipment. If the LHDÆs, trucks, locomotives, etc are automatically guided from origin to destination, the operator is free to operate several vehicles at a time, performing only those tasks which are too difficult for a computer to do by itself.

Jan 1, 1998

Alluvial mining even of shallow deposits is often economically a marginal proposition and causes significant disruption of the land surface through the need for large scale excavation. In-situ leaching using chlorine-based solutions eliminates the need for excavation, the disruption of surface activities, and may greatly reduce processing costs and improve recovery especially of fine gold. A spreadsheet model has been developed to design and analyse the economics of in-situ leaching of alluvial deposits. In-situ leaching of shallow deposits may be achieved in a number of ways including solution infiltration via perimeter ditches or the use of injection and production wells. In most instances deeper alluvials may only be processed using well systems. Two deposits are analysed in this paper. The first consists of one metre of payable wash beneath one metre of overburden. Some thirty fields per year are treated processing 450 000 BCM per year at a cut-off grade of around $4/tonne ($8/rn3 or the gold value equivalent of 0.26 gh or 0.52 g/m3). The principal costs are well drilling and labour. Sensitivity analysis indicates that doubling the throughput halves the cost. The economics of the project are relatively insensitive to reagent costs and consumptions, and halving the recovery doubles the cut-off grade. Shallow deposits treated using the infiltration ditch system may be marginally cheaper and more conveniently managed although this depends on the relative cost of digging the perimeter ditches. In a second example, deep leads with one metre of wash beneath 30 metres of overburden necessitate significant drilling and feature reduced throughputs as a consequence. In the example chosen 30 fields a year are processed for an annual throughput of 65 000 tonnes costed at around $20/tonne (1.4 g/t Au equivalent, $40 /m3). Sensitivity analysis indicates that throughput, and drilling cost per metre are not important but there is a high sensitivity to recovery and to the depth to payable wash.

Jan 1, 1992

By B Veera Reddy

For thick seam extraction, the Blasting Gallery (BG) method is ideal for obtaining the maximum percentage of extraction at fast rates. The continued use of this mining method is threatened by strata control and spontaneous heating problems. The strata control problems are managed, to some extent, by supporting the roof with cable bolting and roof bolts of 2.5 m length and with regular induced blasting. However, a number of BG panels were sealed off prematurely due to spontaneous heating problems, losing 1.12 million tonne of coal. The use of compressed CO2 for controlling mine fires is not uncommon in the coal mining industry, but its use is something new for controlling active spontaneous heating in a BG working panel goaf. This method was attempted at the BG I/8 panel of the Godavarikhani (GDK) 8 Incline, Singareni Collieries Company Limited (SCCL). The spontaneous heating was effectively controlled and the remaining coal in the panel was completely extracted by continuous flushing of carbon dioxide (CO2) into the goaf for 40 days. One thousand and twenty four CO2 cylinders, each of 45 kg at 80 bar pressure were flushed into the goaf at the rate of 25 to 30 cylinders (350 m3 of CO2) per day. A quantity of 35 510 tonne of coal was extracted during the period from 30 July 2002 to 7 September 2002. The main objective of this paper is to present the complete process of controlling the spontaneous heating by CO2 flushing and to report on the study of the behaviour of gases and their analysis. The unique experience of controlling the active spontaneous heating in the working Blasting Gallery I/8 panel goaf at GDK 8 Incline with CO2 flushing (till the completion of the panel) will go a long way in establishing a working procedure for handling fire/spontaneous heating problems in BG panels.

Jan 1, 2005

The recovery of Copper by solvent extraction was arguably the most significant hydrometallurgical development of the 1960's. However, only dilute solutions at relatively high ph. could be treated effectively. ICVAcorga introduced in the 1970's the much stronger salicylaldoxime based reagents, which resulted in a dramatic change in plant configurations. So for example plants which normally use 3 extraction stages and 3 stripping stages could be reduced to 2 + 2 staging or 2 + 1 staging, even for high copper tenors at low ph. Leach liquors up to 40 g A copper are being treated at some locations so that almost any copper bearing liquor can now be purified by solvent extraction. Following a considerable amount of research and development effort ICI/Acorga have made further significant advances in the application of copper solvent extraction. The first of these is in computerised flowsheeting. In conjunction with ICI's Process Technology Department, Acorga have developed the Minchem Computer program, which uses laboratory generated equilibrium data to facilitate optimisation of metallurgical operating parameters. The program can operate in two modes, prediction and assessment so that given suitable extract and strip isotherms plant performance can be predicted for a given set of conditions. Alternatively existing plant performance can be assessed rapidly and accurately, thus providing a valuable quality control tool. Pilot plant work can now be restricted to evaluation of the physical aspects of the SX circuits, and to looking for the unexpected. Secondly using the patented 5-Nonyl Salicylaldoxime in precise admixture with a number of carefully selected strip modifiers, the Acorga M-range has resulted. Reagent selection from these new extractants can now be based on how the reagent can alleviate the remaining problems such as hydrolytic stability, transfer of iron, and organic losses by entrainment which may be related to particular features of the operating plant. Thus the Acorga M-range has enhanced and significantly extended the use of solvent extraction for copper, whilst refinements in the processing of metallurgical data has considerably simplified circuit design and analysis of results.

Jan 1, 1989

By Lynch K. S, Street R. P

Hamersley Iron's pellet plant at Dampier involves ore drying, dry grind- ing in ball mills, disc balling and, indurating by use of the straight grate system. After five years of operation, the plant's output has been increased from the design capacity level of 2 M.T.P.Y. to 2.5 M.T.P.Y. Steps taken in equipment modifica- tion, improvements in process control, and changes in maintenance procedures are described.

Jan 1, 1973

By P M. Ashley, M T. Costelloe

The Mary Kathleen tailings storage facility (TSF) contains 7.1 Mt of uranium mill tailings. This waste repository was rehabilitated in the mid-1980s using a multibarrier dry cover. At the time of mine closure, it was predicted that acid mine drainage (AMD) generation would not eventuate and that uranium, heavy metals and radionuclides would be adsorbed within the tailings and not seep from the TSF into the environment. This study reports on the environmental status of the TSF. Gamma-ray measurements indicate an intact dry cover of the waste repository. However, seepage (~0.5 L s-1) of acid (pH 5.71), saline (0.31 per cent), high conductivity (8.8 mS cm-1), slightly oxygenated (DO 2.6 mg L-1), radioactive waters occurs from the toe of the tailings dam into the former evaporation ponds and local drainage system. Seepage waters are calcium-magnesium-sulfate rich with elevated Fe, Mn, Ni, U and Zn values. Oxygenation of tailings seepages causes the precipitation of Fe and coprecipitation and adsorption of other metals (Pb, U, Y), metalloids (As), rare earth elements (Ce, La) and radionuclides (U-235, U-238, Th-227). In contrast, Ba, Ni, Sr, Mn, P, S, Ca, Mg, Na and K as well as Ra nuclides (Ra-226 and Ra-228) remain in solution until pH neutralisation and evaporation lead to their precipitation in efflorescences and sulfate-rich evaporative sediments further downstream. Dry capping of Mary Kathleen uranium mill tailings has resulted in physical containment of the wastes. The dry cover shows no signs of erosion and radiation levels are within acceptable limits. However, the flow rate of surface seepage from the TSF occurs at a much higher rate than predicted. Also, the neutralising potential and adsorption capacity of the tailings and underlying aquifer were overestimated at the time of mine closure. AMD has developed mobilising metals, metalloids, rare earth elements, alkalis, radionuclides and sulfate from the waste repository into surface water seepages.

Jan 1, 2003

By R Benton

Since commencement of underground production in March 2006 the Kencana K1 orebody has delivered over 0.9 Mt of ore containing 1.18 Moz Au until April 2009. The Kencana deposits are part of the Gosowong goldfield located on the island of Halmahera in eastern Indonesia. The district has produced over 2.3 Moz since mining commenced in 1999 with the first two deposits extracted by open pit methods. The K1 deposit is technically challenging being a vein style deposit with poor to very poor ground conditions and a relatively low overall dip of around 45 degrees to mine grid east. It is mined under geological control by underhand cut and fill methods utilising cemented paste backfill aiming for 100 per cent extraction. With these conditions the operation was mine limited in the initial two years of production resulting in low stockpile levels for mill feed blending and hand-to-mouth mine to mill production. Mill feed rates have increased from 30 t/h to 55 t/h over this period following business improvement projects while mine production has escalated to match this increase as deeper levels of the deposit are accessed. The operation has now become mill limited. Monthly mill gold production from Kencana has ranged up to 50 000 oz with head grades ranging between 8 g/t and 122 g/t Au on a daily basis. As individual grade control assays can extend into the per cent Au range there are challenges in sampling, modelling, mining, blending and milling the Kencana orebody. At the time of commencement of ore driving in K1 the national geology team had little prior experience in underground mining and systems were in early development. The variances in feed grade and tonnages have been reduced with time as the operation evolves and business improvement initiatives are undertaken to smooth peaks and troughs in ore supply. Two such initiatives in the geology department have added over US$20 M of additional revenue per annum.

Jan 1, 2009

The problem of ventilating an advancing drive which finally reached 5,550 ft from the fresh air base in rock with virgin temperature of 83°F was solved economically by the use of two 15 h.p. electric fans forcing ventilating air through 281/2 in. dia. fiexible polythene tubing of wall thickness 0·010 in. The saving in cost over conventional metal duct was estimated to be £3· 24 per ft of drive.INTRODUCTIONAs described by Dutton (1965) and Newton (1965) the existing 1,480' ft level drive at the South Mine, Broken Hill, was extended in two stages. The first extension was 3,790 ft, the face then being 4,790 ft from the fresh air inlet. Later, after a vertical ventilation connection was made 1,000 ft back from the face, the drive was extended a further 4,550 ft. The rate of face advance in both stages was about 15 ft per day (3 shifts of 5 ft), 5 days per week. Drive size was 8 ft by 9 ft 6 in. high.The Mines Inspection Act of New South Wales requires 6 hr shifts (instead of the normal 7 in Broken Hill) to be worked if the wet bulb temperature exceeds 80°F. Virgin rock temperature on the 1,480 ft level was 83°F. Inlet air temperature varied with seasons from 58 to 72,oF wet bulb.To adequately ventilate the face, a flow of air of about 11,000 c.f.m. was thought to be necessary.Re-entry time was about 75 min, being governed by the 7 hr length of shift (3 shifts per day) and travelling time.Blasting was permitted only at the end of the shift, in accordance with local industrial restrictions.

Jan 1, 1965

The Circum-Pacific is defined to include back arcs and continental areas of volcanism. Volcanoes must be regarded as dormant, if the time since their last eruption is less than about ten times the average interval between eruptions, even if they have not erupted for tens of thousands of years or longer. It is important to establish a complete history, over as long a period as possible, for varying volumes of eruptions. In order to assess risk, the average interval between eruptions should be calculated for each range of erupted vo4ume~, in orders of magnitude from about 10 m up to the maximum volume possible at each volcano. Eruptions are considered to consist of individual explosions, and are grouped into eruptive sequences, with the sequences grouped into overall cycles of activity. Frequency of eruption versus volume erupted is a linear relationship for most andesitic and basaltic volcanoes, but is often bimodal at rhyolitic volcanoes, 3 with a slower eruption rate above about 10 km. Worldwide, during the last 7000 years, there have been three known e~uptions with volumes of ejecta of > 100 km , a9d, since 1638 A.D., 7 with volumes > 10 km , and 48 with volumes > 1 km . Average recurrence intervals are about 71 years for eruptions of > 20 km , 39 years for those of > 10 km , and 7 years for those of > 1 km . Rough esti- mates can be made for larger volumes erupted. Mean intervals in the Circum-Pacific are about twice the length of those for the world as a whole. During the last 400 years, about 61% of the world's eruptions of > 1 km3 have taken place in this region.

Jan 1, 1987

For effective backfilling, a system with the 13 features covered under the section: effective backfill systems, is required. Effective backfilling cannot be done with systems that are badly designed, outdated or overextended. But in cases where the requirements for effective backfilling are overlooked, the shortcomings of such systems may wrongly come to be regarded as limitations to the capabilities of backfilling per se. Mining is currently done to depths of 3500 m, and is to go deeper since the gold-bearing reef extends down to 4000 m and 5000 m below surface. Effective backfilling systems will be required at such depths, and with suitable design this can be achieved. Mining the narrow tabular orebody of a typical South African gold mine is described at the start of this paper, then cycloned backfilling is discussed in general, and its effectiveness (most of the fill placed in South Africa is cycloned backfill or ædeslimed tailingsÆ).

Jan 1, 1998

By D Wallace, H Guo, B Shen

Appropriate management of the ground subsidence induced by coal mining activities is essential. To assess the effectiveness of grout injection for mining subsidence control, time-lapse ground penetrating radar (GPR) imaging data were collected before and after grouting. The pregouting subsurface responses were interpreted to have changed after grout injection. The changes were caused by injecting the grout materials (flyash and water) into the delaminated strata and permeable fracture zones. The largest distance of the observed GPR event changes at the right end of the main profile is 132 m away from the injection borehole. Although the GPR images could not easily tell if the image changes were caused by the injected ashes or the water, it does suggest that the injection materials had been pushed at least 130 m away from the injection borehole. This is consistent with the borehole scan monitoring, which confirmed that the grout has travelled more than 90 m from the injection borehole.

Jan 1, 2006

The Vertical Crater Retreat (VCR) mining method was used to mine two rib pillars at New Broken Hill Consolidated Limited (NEHC) mine at Broken Hill. Experience gained in a trial VCR pillar stope was applied to the design of Panel 3CN pillar (P3CN). Restricted access and a central sandfilled plug in this pillar necessitated the use of angled blastholes and blasting of the stope in two sections. Borehole surveys and close charging supervision were necessary to block, charge and stem the spherical charges in the correct position. The laver sections of the VCR stope with rock walls attained an average slice advance of 2.2 m whereas the higher section of the st.ope between cemented fill walls attained an average slice advance of 3.4 m. Achieved fragmentation was extremely fine. External dilution was assessed at less than 10 per cent. A cement consolidated sandfill exposure of approximately 1300 m2 stood vertically with no significant slumping.

Jan 1, 1983

The Beaconsfield Gold Mine within the township of Beaconsfield in northern Tasmania has been established in recent years to exploit a known resource in the Tasmania Reef which was mined to a depth of 450 m below surface between 1877 and 1914 yielding 850 000 ounces of gold. Activity at Beaconsfield over the past 30 years has involved drilling to assess the resource and rehabilitation of the top 100 m of the Hart Shaft that had collapsed since the original closure. More recently the mine has been dewatered to the 375 m Level and work is well advanced establishing infrastructure aimed at recommencing production during 1998 - 99. This paper reviews geotechnical studies undertaken to-date and provides a brief overview of some practical methods of analysing ground conditions. It pursues simple observation and analysis as a tool for assessing mine design and ground support options without recourse to extensive monitoring and measurement and avoids the use of elaborate models before the basics are understood.

Jan 1, 1997

By Doyle J

The direct method of gas content determin ation developed by McCulloch and Diamond of the U.S.B.M. has been adapted to estimate the gas content of coal seams from slim exploration borecores in the Southern Coalfield. Emphasis has been placed on the measurement of this parameter without the physical destruction of the core which is required for subsequent chemical analysis. In field and laboratory equipment is simple and inexpensive. It enables the determination of lost and desorbed gas from a fresh geological core. Residual gas is not measured but is considered to be relatively low and of no consequence in ventilation considerations. Results to date show good correlation with those obtained from underground face samples. Gas contents from the mined seams varying from 2 m3/tonne to 14 m3/tonne have been measur- ed with this apparatus. Gas desorption characteristics of South Coast coals have been found to closely follow exponential mathematical Models.

Jan 1, 1983

By Adams R. W

A commercial scale strong acid leaching plant was commissioned at the Risdon plant of the Electrolytic Zinc Company of Australasia Limited, a member of the Pasminco group of companies, in February 1988. It was designed to upgrade the lead/silver residue product from 26% to 37% lead, thus improving the realised value and increasing overall zinc recovery. This in turn enables the lead smelter treating the lead residue to improve its productivity. The equipment consists of two 250m3 bricklined leach tanks and two 6.5m diameter stainless steel high rate thickeners operating with counter current washing at 95¦C and 120 g/L H2SO4. The present filtration and drying equipment is also being replaced. Two 35m2 squeeze action filter presses equipped with horizontal plates are being installed, together with ancillary drying equipment to ensure that the final product complies with the transportable moisture limit. The new filtration plant is due to be commissioned in mid-1989.

Jan 1, 1989

BHP World Minerals (BHP) commenced development of the Cannington exploration decline in 1993 using Peabody Resources LimitedÆs Mining Services Division (PRL) as the contractor. Approval was given in 1996 to complete the mine development and bring the mine into production in 1997. The total mine development at Cannington by PRL has included 14 km of headings, 1040 m of shaft sinking for ventilation and ore hoisting and 990 m of raise boring for ore passes and vent rises. Hard money schedule of rates contracts were used for mine development from 1993 until early-1997. This took the mine development down to 570 m below the surface and developed six levels. Because of significant risks, the final phase of the mine development was completed in 1997 under a partnering agreement. This paper examines the factors considered in making the change to partnering, the issues involved in negotiating the partnering agreement to complete the Cannington mine development and the outcomes of the agreement.

Jan 1, 1998

The methods and equipment developed and used at the Mount Isa mine to obtain samples of uncemented hydraulic fill lying above major horizontal pillars are described. From existing underground development, samples were obtained at distances up to 50 m, of which up to 30 m was through hard rock. Samples were used to predict fill behaviour as it affected mining security during recovery of the pillars.Specialized equipment for difficult sampling situations and also other sampling exercises are described.

Jan 1, 1976

By A Asadov, D Nair

This paper discusses the determination of the percentage of crystalline material in a natural silica deposit which would be considered harmful to health (<10 ¦m). A sample of the natural silica material was recovered and split into a number of subsamples. Samples were sieved using a conventional dry sieve to remove material >300 ¦m before testing. Particle size analyses of the fine fraction showed 75 per cent of the material was below 10 ¦m. Initially XRD and XPS were run on the samples to determine the minerals and composition of the sample. The XRD revealed strong quartz peaks and smaller tridymite peaks with a high amorphous background. The XPS indicated traces of iron, titanium and carbon with dominant silicon and oxygen peaks. The structure of the sample was investigated using the SEM. It was apparent that there were a number of larger quartz crystals (>30 ¦m) with a mix of amorphous and crystalline in the finer fraction. It was also observed that amorphous material was attached to the surface of the crystals. In order to determine the percentage of crystalline quartz a known amount of quartz was added to the sample incrementally. On each addition an XRD scan was performed and the area under the peaks recorded. Back analysis of these peaks with an assumed percentage of crystalline material in the original sample gave a crystalline content of 30 per cent. The quantity at the respirable fraction was estimated to be 18.7 per cent.

Jan 1, 2003

By M D. Hannington

On the southern flank ofLihir Island, a group of three volcanic cones were discovered at water depths from 1000 - 1500 m. The volcanoes are located in a narrow zone of recent seismic activity and elevated heat flow (up to 100 mW/m2). The recovered rocks consist of fresh alkali-olivine basalts, clinopyroxene-rich basalts, and porphyritic phlogopite basalts. One of the volcanoes has a recently erupted ejecta blanket (recovery of warm ash!), containing a suite of abundant mafic to ultramafic xenoliths. These xenoliths appear to represent a unique cross-section of subduction-modified mantle material sampled by volatile-rich primary fore-arc magma during its ascent to the seafloor. The most spectacular finding was the discovery of an active hydrothermal system with associated vent fauna and mineralization at Edison Seamount, and the recovery of unusually gold-rich (up to 43 ppm Au) hydrothermal precipitates in 1050 m water depth at the top of Conical Seamount, only 25 km south of the world-class epithermal Ladolam gold deposit on Lihir. The discovery of submarine hydrothermal vents associated with highly alkaline, quartz-undersaturated fore-arc volcanics represents a significant departure from similar hydrothermal systems on the mid-ocean ridges and in back-arc basins. The high gold grades in hydrothermal precipitates at Conical Seamount may indicate the first documented example of a shallow-marine epithermal system analogous to those known on land, and has important implications' for the gold potential of submarine arc environments.

Jan 1, 1995