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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

The Kyaukpahto gold deposit is located at latitude 23¦ 47Æ 55" N and longitude 95¦ 56Æ 35" E in the Kawlin-Wuntho district, Sagaing Division, Myanmar. Gold mineralisation is hosted in sandstone of the Male Formation (Eocene) and occurs as stockworks and disseminations. The high-grade gold mineralisation is locally concentrated in a breccia zone. The gold deposit is closely associated with a NNE-SSW trending fracture system that forms tensional open fractures. These tensional zones are thought to be directly related to the Sagiang transfer fault system that experienced several hundred kilometres of northward dextral movement during the post-Upper Miocene. Major sulfide minerals are pyrite and arsenopyrite with minor galena, chalcopyrite and sphalerite. Gold mineralisation is associated with intense silicification, sericitisation and argillic alteration. Primary Type I fluid inclusions in the quartz from the mineralised zone are 5 to 15 ¦m across and yield homogenisation temperatures of 239¦ - 310¦C and salinities of 1.2 - 10.9 wt per cent NaCl equivalent. Laser Raman spectroscopic analysis indicates the presence of CO2 and CH4 in Type II fluid inclusions. Sulfur isotope analysis of disseminated pyrite in the ore zone by the laser ablation method indicates d34S values from -5.48ë to +5.53ë. The geology, alteration, ore mineralogy, fluid inclusion microthermometry and sulfur isotope data of the Kyaukpahto deposit are consistent with an epithermal ore deposit model for the deposit. However, the presence of CO2-liquid bearing fluids and the detectable CH4 in the gaseous phase of the fluid inclusions conflict with a typical epithermal origin for the Kyaukpahto deposit. The host siltstone and mudstone sequence, trace element geochemistry, and the disseminated nature of sulfide and gold mineralisation are also common characteristics of a sedimentary rock-hosted gold deposit, and these features may indicate a Carlin-like affinity. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By Haslam C. O, Taylor S

Four historical periods of Australian lead-zinc-silver exploration and development are identified.In the first period, 1841-1900, Australia's deeply weathered surface yielded bonanza deposits. The outcrops of easily worked secondary ores, enriched in silver and lead, were discovered by prospectors with the incentive of production at times of high silver price. Districts discovered in this period still contributed, in 1982, 60 per cent of Australia's lead-zinc production, mainly from Broken Hill.The second period 1901-1944, was one of strong mine development but stagnant lead-zinc-silver exploration. Mount Isa was discovered and developed, but otherwise effort was devoted mainly to the mining of deposit extensions and to the metallurgical flotation of zinc. The deposit discovery rate was low but the discovery of Mount Isa was so significant that discoveries of this period contributed, in 1982, 30 per cent of Australia's lead-zinc production.The third period 1945-1972 was one of re-invigorated exploration.New geological and geophysical surveys, developing concepts of ore genesis, and worldwide comparisons of orebody settings improved target area selection. Evolving remote sensing methods, and rapid analytical techniques, assisted detection of lead-zinc deposits. Initiation of exploration led to discovery:1. in regions of previously unsuspected lead-zinc-silver potential (Sorby Hills, Mons Cupri, Golden Grove).2. of poorly exposed deposits in recognized mineral districts (Hilton, H.Y.C., Woodlawn, Elura, Que River).3. during active exploration for another commodity (Brown's, Woodcutters, Mary River, Beltana, Lady Loretla and Pegmont).The period saw a sustained growth in lead-zinc-silver production mainly from established operations. Mines discovered during the third period contributed, in 1982, 7 per cent of Australia's lead-zinc production.In the fourth period from 1973 to the present, exploration initiatives became fewer but resulted in continued discovery of poorly exposed deposits at Teutonic Bore, Thalanga, Benambra, Scuddles, Hellyer, and Jillawarra. Development initiatives strengthened in the period and production commenced at Woodlawn, Teutonic Bore, Que River and Elura; pre-feasibility work commenced at Scuddles, Hilton, Lady Loretta and Thalanga.

Jan 1, 1984

EQUIPMENT TO BE DISCUSSED It is the intention of this paper to discuss the transporting of large skid mounted equipment and track machines around open cut mines. By large I mean 100 to 350 tonne capacity being trans- ported on minesite lowloaders and the capability of moving loads of 300 to 1200 tonne by other means. It is also possible to move larger loads. The main factor in transporting large loads is eco- nomics and is a balance between the cost of the transporter and the increased production from the equipment being moved. For example it could be possible to shift a 7000 tonne dragline faster than their own means of propulsion, being the "walking feet" method. However for the cost of $10 to $12,000,000.00 or more you can afford to walk the dragline and suffer the production loss. Let us concentrate on the economically realistic. BOSICH FBT MINESITE LOWLOADERS 100 TO 350 TONNE CAPACITY. Why initially consider having such a transporter? Our experience has shown that initial enquiries evolve from mines requiring to move large drills, 50 to 200 tonne weight and large hydraulic excavators or backhoes which are track mounted. The reason for shifting them is to move equipment from a mining face to enable blasting, to remove to the central workshop for repairs, overhauls and routine maintenance, to transport equip- ment from one mine face to another, and for drills, to move from one drilling pattern to the next. The cost in terms of time lost and extra wear on tracked equip- ment easily justifies the cost of the transporter. The tracks on large drills and shovels are not intended for long distance walk- ing. These tracks are designed to move the machines to position for drill holes, or moving around the mine face. Long distance travel, ie generally 1 km or more, results in overheating of track rollers, track drive components and accelerated wear on tracks.

Jan 1, 1989

The Fiji islands are the world's most ac- cessible Cenozoic remnant arc. They provide occasional subaerial exposures of an arc frag- ment some 750 km long and 80 km wide, or about the size of the NE Japan volcanic arc. The fragment is one part of the former Vitiaz Arc which once connected Fiji to Vanuatu and Tonga. The Fijian portion has the most completely stu- died Tertiary section which contains a history of arc volcanism from the late Eocene to late Miocene (the Early and Mature Arc stages), a rich diversity of subsequent magmatic responses to the most recent episode of arc rifting (Ear- ly Rifting stage), and an example of transition from arc to intraplate volcanism as this rift- ing proceeded (Late Rifting stage). The arc fragment is surrounded by younger backarc ba- sins. The chemical composition of magmas in both the arc and backarc changed with time, but in opposite directions. In the arc, magmas be- came enriched in sialic components from the Early Arc to Early Rifting stages, whereas in the basins the magmas became more similar to mid-ocean ridge basalts (MORBs) with time. In southwest Fiji, the oldest backarc basin has subducted obliquely beneath the remnant arc du- ring the last 3 Ma, resulting in chemically ex- otic calcalkaline volcanism. During the last decade over 150 radiome- tric age measurements and 650 rock analyses, many with trace element and isotopic measure- ments, have been obtained for samples from Fi- ji, Tonga, and the surrounding seafloor (Figs. 1 and 2). These data are summarized, drawing from the publications of Colley and Hindle (1984), Gill et al. (1984), Hawkins and Melchi- or (1985), Kroencke (1984), Malahoff et al. (1984), Scholl et al. (1985) and Whelan et al. (1985), from the specific research papers refe- renced therein, and from the half dozen or so manuscripts currently in preparation by these people and others.

Jan 1, 1987

Shear-flocculation is one of a munber of potential processes for upgrading deposits of finely-grained minerals and, more generally, for separating different types of particles suspended in a fluid.The smaller the particle the more likely it is to follow the fluid streamlines, or be entrained within turbulent eddies, and the less likely it is to deviate from the streamline and attach to bubbles (froth flotation), other particles (selective flocculation) or magnetic poles (wet magnetic separation). As shown in Figure I, a number of ingenious ideas have been proposed to overcome this effect. The suspended particles are assumed to be 1 - 10 µm.Thus, the efficiency of attaching ultrafme particles of the valuable component (solid circles) to air bubbles, as in process A, is poor. The efficiency may be improved by attaching the particles to oil droplets (Ralston, Kart and Newcombe, 1984) to form agglomerates that rise or sink (B), to magnetic particles (C) (Parsonage, 1988) to allow recovery in a magnetic separator (D), to other particles to form aggregates (E) that are then either attached to bubbles (F) or recovered by sedimentation (G) (Warren, 1992), and to pieces of polyethylene (Rubio and Baltar, 1991) that rise or sink (H).In A, C, E, F and H, the forces that are generally used to ensure collisions result in adhesion are hydrophobic forces, hydrodynamic forces and inertial forces (Warren, 1992). These drag particles from the streamlines to allow particle-particle adhesion. In B and E, soluble polymeric molecules have been added to bridge between particles as they approach closely during a collision.The second major consideration is selectivity. Selective adsorption of the hydrophobe or polymer onto the valuable particles is the key. Many subtle effects of surface chemistry are involved in obtaining adequate selectivity. A complication in real systems is that the initial suspension is often naturally coagulated and needs to be dispersed before being selectively flocculated.

Jan 1, 1993

Exploration requires substantial capital, and is an extremely risky business (only 1 in 100 prospects ever becomes a mine). The average lead time, that is the time between discovery and production, varies from around two years for a gold mine, to 15 years for a lead/zinc mine. Since it takes years to find, prove and develop most ore bodies, exploration must be carried out years ahead of the depletion of existing deposits. Investors in exploration are exposed to risk for long periods before knowing whether they will receive a return on the funds invested. The risk, high-cost, uncertainty and long lead times involved in exploration make it an aspect of commerce that most managers and investors find difficult to rationalise or understand. Nevertheless, it is a given that the long-term survival of the mining sector depends on discovery of new mineral deposits, however the immediacy of the need to find or acquire new mineral deposits in order to compete and prosper in the industry is less well understood. This paper examines the commercial realities which underlie exploration for new mineral deposits. The first and foremost reason for exploration is to maintain or increase depleting ore reserves which are the asset base of mining, but there are other compelling commercial reasons for exploring, and these are: to maintain return on investment, to achieve growth or create wealth through discoveries which lead tonew mining ventures, to remain competitive in the industry, to diversify, and to control raw material source for mineral processing and/ormarketing ventures. Exploration is necessary for survival of the industry, and also for survival of organisations competing in the industry. The organisations which discover the best orebodies are in the best position to compete, and mining organisations which fail to find or acquire new mineral deposits rapidly become uncompetitive and unable to maintain return on shareholdersÆ investment.

Jan 1, 1998

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 carbonatite-hosted apatite deposit at the Cajati open pit mine in the Jacupiranga igneous-carbonatite complex is located in south-eastern Brazil. The major styles of apatite mineralisation include veins or veinlets, small pods or patches, disseminated apatite and locally massive aggregates. The metallurgical concentration of the apatite is accomplished by grinding the carbonatite, removal of magnetite with a magnetic drum, size classification with cyclones and flotation to suppress carbonates and collect apatite. To examine the character of the apatite grains and to apply those observations to apatite beneficiation, specimens of carbonatite were collected from recent exposures in the mine and from selected drill holes. This study utilised a variety of microscopy methods, including transmitted and reflected light, cathodoluminescence (CL) and scanning electron microscopy with wavelength-dispersive spectrometers for mineral compositions. It was found that LREE, Na and Sr are the main elements replacing Ca in the apatite, with relatively low oxide contents (<0.5 per cent, but up to about one per cent for Sr). LREE and Sr contents consistently increase in the latest intrusions (C4, C5 and SP24). Chemical and morphological modifications of apatite were recognised to be related to detrimental effects on flotation recovery. Three examples illustrate this relationship: 1.  Apatite with reddish CL is fractured. Locally that apatite exhibits no CL due to CL quenching caused by iron entering the apatite structure.2.  Apatite from the xenolith-rich area exhibits abundant fine, platy silicates that appear to interfere with flotation and lower apatite recovery.3.  So called æhardÆ apatite, which is more difficult to grind, occurs in an area in the mine that was affected by hydrothermal activity. That activity is evidenced by the development of a bright, white CL rim in the apatite, and by the presence of late veins of carbonates and sulfides that fill fractures and appear to contribute to higher cohesiveness of those apatite grains.

Jan 1, 2008

The Drummond Basin is a large intracratonic basin which developed in central Queensland during Late Devonian to Early Carboniferous times. It is postulated that the Drummond Basin developed as a foreland basin, to the west of a probable west dipping continental margin volcanic arc and subduction zone. The Drummond Basin sequence has been s,ibdivided into three tectono-sedirnentary cycles. Cycle I forms the base of the sequence and comprises thick sequences of intermediate to felsic volcanics and proximal volcanolithic sediments. The volcanics, particularly in the northeast of the basin, occur in separate sub-basins around what were probably isolated eruptive centres. In contrast, the rocks of Cycle 2 and Cycle 3 have been deposited during episodes of broad, basin wide subsidence. Five major igneous episodes, some of which may be co-magmatic, are recorded in the eastern Drummond Basin and environs between the Late Devonian and the Early Permian. The first and second episodes occurred during the Late Devonian to Early Carboniferous and are represented by the volcanics in Cycle 1 and Cycle 3. The third episode involves the intrusion of granitoids of Early Carboniferous age into the Drummond Basin sequence. These occur in the Ukalunda area and are related to gold mineralisation at Ukalunda and Valhalla. Episode four includes the felsic volcanics of the Bulgonunna Volcanics. This episode does not appear to he broadly auriferous but gold mineralisation does occur in possible basal Bulgonunna Volcanics at Mount Coolon. The final episode includes a suite of granitic intrusives which intrude the Bulgonunna Volcanics and which are considered to be co-magmatic with them. These intrusives do not appear to be generally associated with gold mineralisation. Gold mineralisation is widespread in the north- eastern Drummond Basin and environs. It is probable that the gold mineralisation is concentrated in epithermal cells during the Late Devonian to Early Carboniferous volcanic cycles. This style of mineralisation appears to be related to tectonic setting in modern analogues.

Jan 1, 1989

By G Harbort, T Georgiev, M Phillips

Phu Bia Mining Ltd, a subsidiary of Pan Australian Resources Ltd owns and operates the Phu Kham copper concentrator located in the Xaysomboun District, Lao PeopleÆs Democratic Republic. Construction and commissioning of the 12 Mt/a copper concentrator and associated facilities was undertaken by Ausenco Ltd and handed over to Phu Bia Mining Ltd in April, 2008. The first copper concentrate was produced in April 2008 and the ship carrying the first shipment of copper concentrate set sail 27 July 2008. Nameplate mill throughput was reached in August, 2008 when 1 020 000 dmt of ore was treated in the milling and flotation circuit. The Cu-Au mineralisation is contained in skarn and stockwork mineralisation and is interpreted as a distal expression of a conventional porphyry copper-gold system. Copper mineralisation is principally in the form of chalcopyrite with lesser bornite. Gold occurs as small grains, typically associated with copper minerals. In the upper portions of the orebody, oxidation and water table movements have resulted in near surface leaching of the copper and below the leached zone, supergene enrichment resulted in chalcocite mineralisation. The circuit operates with a semi-autogenous grinding (SAG) û ball mill circuit, feeding a bank of OK200 rougher cells. Rougher concentrate is reground in an IsaMill prior to entering a cleaner circuit consisting of three stages of cleaning and one stage of cleaner scavenging. The concentrate typically assaying 24 per cent Cu and 6 - 7 g/t Au is hauled for over 1000 km across the border to the Port of Sriracha in Thailand, and stocked in sheds at the wharf from which it is shipped to various Asian smelters. This paper discusses the successful commissioning and ramp up to design tonnage and metallurgical performance. In addition, a program to optimise both the comminution and flotation circuit is detailed with reference to grinding efficiency, flotation residence time and size by size mineral recovery.

Jan 1, 2009

The care necessary to ensure a reliable estimate of average value from the individual values of a series of bore holes sunk in any alluvial deposit containing coarse gold was well exemplified in checking over the mining returns from a previously bored deposit in Central Otago, N.Z. While the working returns are always liable to vary considerably from bore values in any particular sub-section of an area, still, over any considerable area, the two must check within small limits, if profits are to be calculated and capital raised on the average value given by boring.The present figures have probably more value than those from, dredge returns since accurate measurements are difficult to obtain in irregular water-filled dredge paddocks, whereas in the case under review the paddock walls stood almost perpendicular, and, as the measurements were taken and checked in the empty rectangular excavation, the yardage estimate may be taken as substantially exact.NATURE OF DEPOSITThe area consists of a narrow valley deposit of fluvio-glacial origin, the original deposit having been eroded and the gold reconcentrated by post glacial rivers. The wash is between 80 and 90 ft. deep, and is formed in two distinct layers:(i.) The top layer from surface to 70 ft. or 80 ft. consists of about equal quantities of gravel mixed with glacial drift, and a small amount of clay. The gravel is typical discoid glacial stone of a maximum size of about 4 in. diam. by 1 in. thick, and consists of soft micaceous schist and laminated slate. This upper layer of wash contains no coarse gold whatever, and very little fine; indeed, in mining, the top 30 ft. is discarded without being passed over the boxes.(ii.) The lower layer of wash to bedrock is similar to the upper, except iihat it contains about 10% of larger discoid stone up to 12 in. diam. by It in. thick; 90% of the gold is concentrated in this bottom layer.

Jan 1, 1929

By M Csar

The Nicolsons Find gold project is located 35 km southwest of Halls Creek in the east Kimberley. It was discovered in the late 1980s by John Nicolson and Stuart Skoglund, when high grade gold mineralisation within outcropping gossanous quartz reefs was found. The area has produced a modest 28 150 ounces of gold over two brief mining periods in 1996 and 2002 - 2003. Indicated and inferred resources as of September 2011 stand at 1 384 000 t at 5.58 g/t Au for 248 500 ounces of gold.The project is situated within the Halls Creek Orogen. Principal host units are metasediments and volcaniclastic sediments of the Koongie Park Formation. Minor mafic units are also recognised. Immediately east of these units, emplacement of the loadstone granite is believed to have provided the structural setting for the later gold mineralising event.Mineralisation is hosted within a broad, brittle-ductile arcuate shear zone about the western contact of the Loadstone granite. In the Nicolsons Find area, the shear zone strikes north-northeast and within the shear zone, gold is hosted in a series of narrow, quartz rich veins up to 200 m in strike length, plunging moderately to the north-northeast. To date, mineralisation has been drilled to a depth of 360 m below the surface. The veins are stepped or en-echelon in nature with low gold values towards the quartz vein limits where compression is inferred and higher grades present in the dilational body of the quartz veins. Gold is fine grained and can be present with or without associated sulfides. Sulfides are generally minor and consist mainly of arsenopyrite with traces of chalcopyrite, pyrite and sphalerite. Where present, sulfides are generally indicative of higher gold values.CITATION:Csar, M, 2012. NicolsonÆs Find gold project, East Kimberley, Western Australia, in Proceedings Narrow Vein Mining 2012 , pp 31-36 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 26, 2012

By Wang S

Fragmentation is the basic concern in rock blasting and serves as the main measure of blasting effectiveness. A good estimation of fragment size distribution before blasting is very useful and helpful for selecting the proper equipment for loading, transporting, and crushing rocks. A fragmentation model is presented that is based on a hole by hole analysis and correlates the rock properties, burden, spacing, and explosive column with the blasted volume and fragmentation obtained. The drill bench is divided into three separate areas corresponding to different breakage mechanisms. They are: 1) a main breakage or casting area, 2) a secondary breaking area, and 3) an overbreak area. The boundary of the main breakage area is found from the magnitudes of strains in the stress waves and the tensile strength of the rock. The secondary breaking area is determined by stress wave reflections at discontinuities and interactions between colliding stress waves from adjacent boreholes. The overbreak area is that beyond the main breakage area "where the reflected tensile waves no longer break the rock. This analysis is applied on a hole by hole basis to locate the breaking points. Then connecting the breaking points on each side of a borehole forms the triangle of the main breakage area. This is repeated for the next hole and the remaining area is the secondary breakage area. After the angle and area for each hole are found the Kuz-Ram model is used to predict the fragmentation size distribution for each hole. Then summing up the size distribution for each hole gives the total fragmentation distribution. This analysis has been successfully applied to field tests performed in a limestone bench where the entire spoil pile was screened to determine the fragment size distribution. Knowing the explosive and rock properties the breaking area was found for each hole, the strain and stress values, the breaking point, and angle of breakage (104 degrees). The fragmentation size distribution was then calculated and compared to the field data with excellent agreement.

Jan 1, 1990

By Hemingway J, Read T. A, Anderson V. G

IN the determination of very small amounts of lead, the usual analytical methods are far from being satisfactory. Small amounts of lead may be present in the reagents empioyed, and it is customary to run a blank experiment to correct the error arising in this way. When the correction is of small magnitude compared with the amount of lead in the sample, this procedure may be adopted; but when very small amounts of lead are being determined it may happen that the lead found in the blank is as much, or even more, than the lead actually present in the sample, and the analysis has to be rejected. Moreover, the blank experiment cannot check errors which are not common to the blank and the analysis itself eg, the apparatus used in the parallel experiments, etc. It seemed to be desirable, tl;l.erefore, to make an investigation of the problem to ascertain, as far as possible, the nature of the errors arising from the introduction or loss of lead during the analysis, and how far these could be eliminated or corrected for; also, to determine the limits of accuracy of the improved method and the minimum amount of lead that could be reliably estimated.In the investigation, reliability and sensitiveness were regarded as essential features. Speed of operation, on the other hand, was considered of less importance. The amounts of lead dealt with being usually less than 1 mg, there is no alternative to the use of a colorimetric method. The latter, depending as it does upon the development of a brown coloration when ammonium sulphide is added to a very dilute -solution of a lead salt, is far from being a characteristic test for lead. Indeed, it is well known that the salts of most of the commonly occurring metals interfere with this test. Hence it is of vital importance that the lead should be isolated in the form of a pure compound possessing quite definite and characteristic properties before the colorimetric estimation is proceeded with...

Jan 1, 1921

By B L. Dickson, J H. Hodgkinson

The self-organising map (SOM) is an exploratory data mining technique that is both non-traditional and underutilised. Methodologies based on SOM tend to be data-driven and unsupervised, which makes them ideal to assist in the integrated analysis and interpretation of complex and disparate æmineral explorationÆ data sets. While traditional statistical multivariate approaches have difficulty with relationships that are non-linear and data distributions that are not normal, SOM-based data mining procedures are useful in these circumstances. In a SOM analysis each sample is treated as a vector in a data space defined by the variables; and measures of vector similarity, such as the dot-product or Euclidean distance, are used to order or segment a data set into naturally occurring populations. These groupings are positioned as nodes, or groups of nodes, on a 2D rectilinear representation of the multi-dimensional ædata spaceÆ, which is the æself organized mapÆ. While it is common not to include a sampleÆs locational information in the actual SOM analysis, such information can be used to display the spatial location of samples coded by their SOM node or cluster. If one finds there are coherent spatial patterns belonging to samples from particular nodes, or groups of nodes, then there is strong evidence that the technique is determining patterns and relationships within the data that have natural significance. Variations of the SOM technique may be used to perform broad categories of operations, including: 1. outlier identification û for samples belonging to various nodes or grouping of nodes; 2. function fitting, prediction, estimation or imputation û for calibrating relationships between variables and determining replacement values for missing, null or censored values; 3. clustering, pattern recognition or noise reduction û for determining the natural patterns within a data set; and 4. classification û for classifying samples into the clusters determined in three. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By Fredley D. C

Georgius Agricola wrote in De Re Metallica in 1556 that a prudent a miner should not work in a region ruled by a tyrant-a person inimical to mining. In 1926, Associate Chief Sherman of the United States Forest Service in a speech stated, "It is the policy of the Forest Service to encourage the development of the mineral resources within the boundaries of the National Forests ... a man who fails to cultivate a friendly, helpful, cooperative spirit in his relations with those who are endeavoring to develop the mineral wealth of our mountains has no place in the Forest Service." Both statements are as true today as they were in 1556 and 1926. The challenge to the Forest Service and the mineral industry today is to maintain that cooperation in a world that has changed significantly since Agricola and Sherman. The United States Forest Service manages 77,000,000 hectares of land-mostly in the Western United States-an area equal in size to New South Wales. The vast majority-75 million hectares are National Forests, and 1.6 million hectares are National Grasslands. National Wilderness Areas account for 13 million hectares or rough- ly 17 per cent of the National Forest System. National Forest System lands (NFS) are some of the most visible, scenic and controversial lands in the country. They are also some of the most valuable for their mineral potential and production. The value of mineral produced exceeds US$4 billion per year. There are over 1200 producing mines, 15,000 oil and gas leases, and some of the largest lead, copper, silver, gold, platinum, molyb- denum, and coal deposits in the hemisphere. Every year the Forest Service approves nearly 5,000 mineral operations on the NFS. Much of the NFS (75 per cent) is open to location of mining claims pursuant to the 1872 Mining Law. And an equal amount of land is legally available for oil and gas leasing.

Jan 1, 1991

The following is a brief summary of the Bulli Colliery and its method of operation when Australian Iron and Steel Ltd. took control.Bulli Colliery is situated about a mile west of the village of Bulli and is some 10 mUes north of the Australian Iron and Steel Company's Steelworks at Port Kembla.The mine is owned by Australian Iron and Steel Limited, who purchased it in 1936 from the liquidators of Bulli Colliery and Coke Works.It is one of the oldest working coal mines in Australia, having been opened about 1858 and has worked more or less continuously since that date.The seam worked is the No. 1, or Bulli, Seam, which outcrops along the escarpment of the Illawarra Range and is generally operated by tunnels, the portal at Bulli being some 440ft. above sea level.The colliery holding is fairly extensive, but the coal seam over a large portion of the area within two miles of the outcrop has been more or less destroyed by an intrusion of basalt.At present there is no coal being worked within a distance of about 21 miles from the surface, the farthest workings being about four miles from the tunnel mouth.The existing main haulage road, for the first two miles from the portal, is not straight; it is very narrow over fairly long stretches and has many changes of grade varying from 3% in favour of the load to 13% against the load.The dumping station and screens are situated close to the portal of the haulage tunnel, and at this point the coal is loaded into railway hopper waggons, which are then hauled by rope to the top of a self-acting incline. The waggons are lowered down the incline, a distance of approximately! mile, the ruling grade being about 18 %.From the bottom of the incline, waggons are hauled by Company's steam locomotive to exchange sidings from whence they are taken by New South Wales Government Railway engine to the Company's Steelworks at Port Kembla.

Jan 1, 1947

By G Power

The successful implementation of sublevel caving (SLC) at Ridgeway Gold Mine (Ridgeway) requires the technical aspects of caving and flow behaviour to be understood. Cave propagation must be controlled and continuous to allow safe extraction of ore below. The understanding and use of processes to manage granular flow behaviour are also required to ensure economic metal resource recovery. Caving and flow technologies are thus critical to the economic viability of the application of the SLC mining technique to Ridgeway Gold Mine. At the time the Ridgeway SLC design was approved there were significant technical uncertainties associated with cave mining methods, including the SLC mining method in competent rock masses. This was in spite of the fact that caving is a relatively low cost mining method. Unique features generating technical risks at Ridgeway include: the fact that the orebody is blind (550 m below surface); highly competent cap rock (RMR >50); and high aspect ratio (height to width ratio of 3:1). Due to the presence of these unique features Ridgeway is considered to be beyond the current state-of-the-art in SLC mining technologies. The success of the technique at Ridgeway to date has indicated that blind SLC mines in strong rock mass can be operated safely and productively, if the technical risks are understood and properly managed. The risk management process was defined by the use of a formalise process to identify and assess the risks to the mine. The major mining related risks were formalised into Major Hazard Management Plans (MHMP) which contained a description of the risk, early triggers indicating the creation of a hazard and a series of agreed responses to be automatically initiated in the event of each trigger occurring. This paper outlines the following with respect to the critical technology areas of caving and flow: international industry, state-of-the-art technology; gaps in the application of this technology; the application of risk management, new processes developed at Ridgeway to bridge these gaps, including issues encountered; and trials and experimentation conducted to develop the new processes.

Jan 1, 2003

A method is presented for calculating the number of cycles of stress to failure of a drill rod from operating variables; a mathematical analysis by Cooper of the stresses present in drill steels is applied to the forged collar, i1i-in. hexagonal hollow drill steel in common use in Australia; and an estimation is made of the dry fatigue limits at various points along a drill rod.It is shown that these calculations predict failure at a 'point where failure rarely occurs, and do not predict that failure should occur where it does in practice. The reasons for this anomaly lie in the neglect of bending stresses in the stress calculations, and in failuredue to corrosion fatigue at a preferred section rather than to simpler fatigue under non-corrosive conditions.INTRODUCTIONIt has long been recognized that failure of metal sections by fatigue is influenced both by the number of cycles of stress imparted to the section (N) and by the magnitude of this cyclic stress (8). In this paper, both Nand Shave been calculated for rods which have failed in service, and whose operating history is known. While the calculation of the number of cycles to failure is·simple, the calculation of stress is more complex. The results of stress caIculations predict that failure should occur only hi short rods, and then preferentially at the tip end rather than the shank end.However, in actual fatigue failures of steel rock drill rods, approximately 75 per cent. of such failures are close to the shank end, and failures near the tip end are rare.The usual explanation given for this high incidence of shank-end failures is that at the point of failure there exists, in forged collar steels of the type common in Australia "hardness valley" or. "m'etallurgical notch". This narrow valley in hardness is a weakness, and failure is expected to occur at the weakest point at the bottom of...

Jan 1, 1955