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By Arumba J

In the Mount Victor area of the Eastern Highlands Province, re- connaissance geological mapping and new micropalaeontological data have revealed the existence of at least three small thrust- bound windows exposing late Middle to Late Cretaceous (Late Aptian-Early Albian and Late Santonian-Maastrichtian) sedimen- tary rocks and pillow lavas, Latest Paleocene- Earliest Eocene pillow lavas, and Latest Paleocene-Earliest Eocene bathyal micrites. Late Middle Eocene to Late Eocene inner neritic bioclastic micrite units also occurs in thrust contact with Middle Miocene Yaveufa Formation volcanolithic conglomerates. The Cretaceous and Early Tertiary units appear to be thrust-bound and are best exposed in a window just south of Mount Victor, and in two smaller windows to the east. The inner neritic Eocene limestone occurs in a thrust-bound sliver within the Yaveufa For- mation near Obura. The windows occur within extensive out- crop of Early Miocene Omaura Formation and the Yaveufa For- mation. Thin-bedded grey mudstone, siltstone and sandstone, typical of much of the Chim Formation form only a small part of the exposure of Cretaceous rocks. Thick-bedded volcanolithic conglomerate, spilitic pillow basalts, and welded basic tuffs domi- nate the Cretaceous section. The stratigraphy and structure of the Mount Victor area is similar to the Menyamya area, the Aure Scarp, and perhaps the Tapini area. This study has raised the possibility of widespread thrust- ing in the Kainantu area, and suggests that the boundary between the Omaura and Yaveufa Formations is thrusted, at least locally.

Jan 1, 1997

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 M P. Sandy, J B. Simpson

The empirical stope design techniques developed by Mathews et al (1981) and refined by Potvin (1988) and others, have gained rapid acceptance in Australia in the 1990s as a simple, æfirst-passÆ means of designing primary stopes. However, it is recognised that these æstability graphÆ methods can only provide broad design guidelines. Recently published stability graphs (Hutchinson and Diederichs, 1996) encompass a wide range of geological environments, which are not discussed, and the influence of time and blasting practices cannot be easily reviewed. The æscatterÆ in the graphs must reflect, to some degree, different operating practices as well as subtle but important differences in the pre-mining conditions that are not adequately accounted for in the æmodified stability numberÆ. Development of a ælocalÆ, site-specific stability graph may overcome some of these problems. At NormandyÆs Golden Grove Operations, Western Australia, detailed reviews of stope stability performance, including Cavity Monitoring System (CMS) surveys, are undertaken for every stope, with the objective of developing a series of local stability graphs, customised to site conditions and calibrated in terms of æAverage Falloff ThicknessÆ (AFT). Cablebolting of stope hangingwalls is used to develop reinforced æbeamsÆ on each sublevel, effectively dividing the full span into smaller, more stable æsubspansÆ. By analysing stope stability performance in terms of fall-off, cost-benefit studies can be undertaken of the effect of cablebolting and other options for modifying stope spans. The approach used to develop the local stability graphs, and their integration into routine stope design procedures at Golden Grove, is described. The experience gained during stope extraction in the Scuddles orebodies has been successfully applied to the Gossan Hill æC-PanelÆ orebody. This methodology should be applicable to other underground mining operations.

Jan 1, 2000

By Ross J. F

A pillar 'is a block of coal left in the first working of a coal seam for the purpose of supporting the superincumbent strata in such a manner that the winning out roadways will be safe to traverse by men and by the coal transport system.PRINCIPLES OF PILLAR EXTRACTION.The size of a pillar is chiefly dependent on the depth of cover, and in a lesser degree on the nature of the coal, or the condition in which the seam has been laid down. In a seam at 1,100ft. depth of cover, and assuming a spec'ific gravity of 2.5 overall for the cover strata, a pressure of 1,200lb. per sq. in. is the loading on the top surface of the coal seam before mining begins.When mining commences, and as the pillars are formed, an increased load is imposed on the coal left in the pillars, according to the percentage of coal extracted in the first working. Thus, if 25% extraction is taken in the first working, the load on the pillars is theoretically increased from 1,200lb. per sq. in. to 1,600lb. per sq. in. over the area mined. If this state of things went on indefinitely it would not be possible wholly to extract a pillar because of the extreme pressure tliat would be thrown on the diminishing portion; but, fortunately, another element, that of time, enters into and assists the process of extraction.If the superimposed strata were of an entirely plastic nature, this time factor could not operate, as the cover weight would settle instantly on the remaining portion of the p'illar and finally overwhelm it. The cover strata above a pillar generally have a certain rigidity which does not allow the superimposed weight to settle instantly on the pillar fraction, and it requires a period of time to elapse before the extra weight is felt by the pillar.The time factor varies widely, and is governed entirely by local conditions; it can only be determined by experience gained on the spot or in close proximity to the area of operations. It is dependent on the strength of the strata and its characteristics under bridging conditions.It is, therefore, logical to assume that the more rapidly a pillar can be extracted the greater will be the safety factor and also the percentage of recovery of coal.Up to the present time, the operation of the extraction of pillars in this country has been, with a few exceptions, a manual one, and before mechanization of Collieries can be fully accomplished, the process will have to be performed mecharrically.

Jan 1, 1947

Australia had an active uranium mining industry in the 1950s with uranium oxide being supplied under contract to the atomic energy authorities of the U.K. and U.S.A. In the 1960s, when these con- tracts concluded, domestic production ceased, apart from Rum Jungle, whose production until 1971 was stockpiled. By the 1970s, the demand for uranium from the electrical utility industry had risen sharply and the effects on Australia were notable. The Mary Kath- leen mine re-opened in 1976 but, more significantly, exploration activity suddenly expanded with the re- sult that several world-standard deposits were locat- ed. Western world requirements for uranium based on increasing energy consumption and a changing energy mix, will warrant the development of Aust- ralia's resources. By 1985 our mines could be pro- ducing 9,500 tonnes of uranium oxide yearly and by 1995 the export value from uranium could reach that from wool. In terms of benefit to the community the economic rewards are considerable but, in terms of providing energy to the world, Australia's uranium is vital.

Jan 1, 1978

By D R. Rankin, J J. Fuykschot

Forrestania Nickel Mines is owned and operated by Outokumpu Mining Australia Pty Ltd and is situated 400 km east of Perth, WA in the Forrestania Greenstone Belt. Production started in late-1992 at a rate of 600 ktpa and reserves were exhausted in August 1999. Cosmic Boy is the largest of three mines worked and operates in a high stress environment with both very high and very low rock mass strength domains. Stoping commenced in 1995 using conventional benching with long hole slot rises but because of drilling and blasting problems the mine switched to airleg rising. In 1997 the mining method was changed to smaller scale up hole retreat benching with each stope requiring a slot rise of up to 16 m length. This method does not allow top access to the rise location and so airleg rising was used until mid-1998. At this stage mining was approaching areas of higher stress for which it was clear that an alternative to this entry method of rising was required. Small diameter raise boring and Cubex rigs were considered but because of the set up costs were not considered to be viable. Since the ore is relatively soft, burn cuts were developed to make use of specially designed 152, 203 and 254 mm diameter reamer bits (depending on the rise length) and drilled with conventional long hole drilling equipment. These rises are fired in one blast together with the slot rings. In the northern lower section of the mine (below 400 m) the principle stress has been estimated at 60 MPa and this causes minor strain bursting in BIF development headings and loose material to form on walls and backs. Trials of destress blasting modelled on Canadian experience has greatly reduced strain bursting at the face. However only around 300 m of development was affected and so this method had not been optimised when mine development was completed.

Jan 1, 1999

By Wall J, Platell R, Knight P

The Argyle diamond mine is located in the East Kimberley area of Western Australia, some 100 km south of the town of Kununurra. The diamondiferous Argyle olivine-lamproite diatreme was discovered in August 1979. Production began in December 1985 at a planned throughput of 3 Mtpa to recover 25 million carats of diamonds. Presently, Argyle is the world's largest producer of natural diamonds and produced over 40 million carats in 1993 from seven million tonnes of lamproite ore treated. Two phases of evaluation have been carried out. The first in 1981-82 was designed to prove around 60 million tonnes of ore to justify the establishment of a mine with a minimum life of 20 years. The second phase, in 1989-90 involved drilling the orebody to greater depths to provide data on which to plan the final pit design. The final pit contained 120 million tonnes of ore at 5.3 carats per tonne (ct/t) grade. Mining of the planned pit will extend the mine life to around the year 2004, following which it is possible that a further cut-back could be mined. There is also the potential for future underground mining. A substantial amount of geotechnical work has been undertaken due to the fact that the pit will be nearly 500 m deep and is located in a mobile zone causing abnormally high residual stresses to occur in the pit walls. The occurrence of low strength rocks at depth in the pit walls further complicated the situation. Hydrogeological investigations were done to determine pit dewatering requirements and to assess the effect of dewatering on pit wall stability.

Jan 1, 1994

By F Cappuccitti

A colourimetric technique, originally developed for methyl iso butyl carbinol (MIBC) based on the Komarowsky reaction, has been extended to a wide range of common mineral flotation frothers. The procedure, which is briefly described, is shown suited to on-site determinations in a campaign at Kennecott Utah Copper Copperton Concentrator (Rio Tinto). Some observations are described, including: frother (MIBC) concentration evolution down a bank, frother distribution between pulp and froth phases, and loss of frother over time due to adsorption on certain minerals. A possible interference with a fuel oil addition was overcome. MIBC in recycle water was also measured. The ability to monitor frother easily will help interpret gas dispersion behaviour and should permit further optimisation of flotation cell hydrodynamics and chemistry. The capability to directly measure should help optimise the type, amount and location of frother addition, and to assess frother recycled with reclaim process water.

Jan 1, 2005

Most of the gold reef mined today in South Africa is approximately 1 m in width with slopes about 30 m long. To prevent production interruptions it is essential that the stopes should not be blasted in sections. The shortage of experienced labour and arduous working conditions at the face dictate the need for simple blasting techniques.Hence, a pyrotechnic blasting system of sequential firing is in use at present, although research work to improve this technique is in progress.

Jan 1, 1985

By Hengst B

The first production stage in any mining operation is the breaking of whole rock into fragments small enough for loading and subseq- uent handling. This primary fragmentation is most frequently accomplished with blasting. The maximum fragment size and the fragment size distribution have an influence on the subsequent loading, handling and crushing costs. It is assumed that total rock fragmen- tation in practice is a result of an interact- ing array of radial cracks from various blastholes. The representation for the basic pattern of radial cracks emitted from a particular blasthole is first described. A model is then developed simulating the interaction between blastholes at various burdens and spacings, and in simultaneous and delayed modes. The resultant picture is analysed and a fragment size distribution calculated. Methods of incorporating the existing faults in the rock are also discussed. Finally application of this model to various facets of mining are considered.

Jan 1, 1977

By Brzezniak W

Conversion to trolley assist is a means of minimizing diesel fuel consumption, which in many mining situations offers a rapid payback of capital investment plus the inherent flexi- bility and maneuverability of truck shovel op- eration. Whendiesel electric truck shovel mining opera- tion with trolley assist is compared to alter- native mining methods such as conveyor systems, the trolley assist system can be easily justi- fied in terms of return on investment, payback and increased productivity. A review of present technology, performance requirements and past experience in the min- ing industry has led to the development of a trolley assist system for diesel electric haulage trucks which, is proven in the mining environment, offers significant reduction in fuel consumption and increased productivity.

Jan 1, 1981

The Nelson province was chosen as the prime area for rhenium sampling, based on its mineral associations. Sample preparation involved benefication by grinding and flotation. A modification of published procedures for rhenium analysis was developed and the results compared with those obtained for molybdenum on the same samples using atomic absorption spectroscopy. The results of the analysis for both metals is discussed and the relationships between rhenium and the molybdenum mineralisation of this region is commented on. A metallogenic province is defined for molybdenum and rhenium and the tectonic setting for this is discussed. Some consideration has been given to rhenium possibilities in Fiordland and the Coromandel Range.

Jan 1, 1980

By Kerr M. H, Slaughter P. J

Design of the new cemented hydraulic fill plant was based mainly on experience with a succession of previous fill plants. Two stage cycloning of combined concentrator tailings was introduced. Design of the two stage cyclone configuration was assisted by computer simulation of cyclone performance. Cementing additives, namely Portland cement and ground copper reverberatory furnace slag, drawn from bulk storage, are pre-slurried prior to mixing with hydraulic fill. Quality control of fill being sent under- ground is achieved by a series of automatic control loops. Plant instrumentation was given prime consideration during the design phase. Inventory checks of all mass flow metering equipment are made at regular intervals. Fill is pumped to boreholes at several locations. Details of the equipment and operating procedures for the plant are Discussed.

Jan 1, 1973

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

Builders have been very familiar with iron since ancient times, but because of its weathering characteristics it was relegated to use as a fastening material in unexposed places and bronze became the popular metal for exposed applications. It is of interest that iron nails buried deeply about A.D. 87 at a Roman fortress and carefully covered with clean beaten earth under conditions practically excluding oxygen were found to be almost unrusted 1870 years later. (1) Today, weathering steels, which contain less than 2% of alloying additions, can be used in exposed situations without protection and without experiencing continuous wastage. As we all know the steady access of atmospheric oxygen to the surface of plain steel causes continuous corrosion. Weathering steels, on the other hand, due to their alloying additions develop their own impervious surface layer which limits the access of atmospheric oxygen and effectively limits corrosion. Table 1 shows the chemical composition of a plain carbon steel and a typical weathering steel. Fig. 1 shows what we refer to as direct corrosion - time curves in which corrosion loss is plotted against time of exposure in an strial environment. The weathering steel has a high rate initially which then decreases steadily for about three years, as the stable rust film builds up. This is followed by a practically linear relation where the very low corrosion rate, according to one school (2), is due to a blocking of the rust pores by insoluble complex salts formed from interaction of the oxidation products of the separate alloying elements. The start of the linear region represents a total corrosion loss of only about 2 thousandths of an inch.

Jan 1, 1970

Shortly after the outbreak of war in September, 1939, it became apparent that considerable quantities of certain metals and alloys which at that time were not being produced in Australia would be required in increasing amounts for the prosecution of the war effort.Australian industry had relied on overseas sources of supply for the relatively small quantities of many special metals which were needed under peacetime conditions. As the war progressed so great became the demand for mooy of these alloys in Great Britain and other allied countries that it was necessary for them to place an embargo on the export of these materials, and thus it soon became evident that if Australia's production of munitions was to expand and develop it would be necessary for her to become independent of overseas sources for many of these essential 'items. The shipping position also became very acute, and this, accentuated the difficulties of obtaining these alloys. Because of the conditions outlined aqove, the manufacture of over thirty different metals and alloys .was undertaken at Newcastle during the past decade. It is with two of these that this paper proposes to deal. They are tungsten powder and ferro-silicon.TUNGSTEN POWDERTungsten is a rare and comparatively new metal. It is used extensively in the production of high speed tool steel, some grades of which contain up to 22% tungsten. Such articles as twist drills, reamers, milling cutters, and lathe tools are made from tool steel, and the demand for these during the war years was greater than ever before. The use of tungsten in the manufacture of high speed steel was first developed in 1914 by the Germans, and when it was discovered that they were using tungsten bearing steel for the production of munitions there began a world-wide search for ores of this rare metal.

Jan 1, 1947

The HYC mineral deposit at McArthur River is about 1.5 kilometres long by 1.0 kilometres wide and 55 metres thick, and contains 227 000 000 tonnes of ore of average grade 9.2% Zn, 4.1% Pb and 41 g/t Ag. Major studies have been carried out continuously since 1955 over the deposit and in the general region. This work is continuing. A small underground mine was developed by driving a decline tunnel to intersect the mineralization 80m below the surface. This provided a fully representative sample of the total deposit unaffected by weathering or oxidation, and was available for the major test program carried out in a pilot plant built to determine the full metallurgical charac- teristics of the ore. The pilot plant operations followed investigations by research institutions, both in Australia and overseas, into the problems of this ultra-fine grained ore. The ultra-fine grained nature of the ore required grinding to an average size of 95% passing 37 micro metres or 400 Tyler mesh. A major feasibility and environmental report was submitted to the government covering all aspects of a development program. The geological, engineering, metal- lurgical, economic, cultural and environmental submissions contained in this report were studied by external consult- ants employed by the government, who agreed with the conclusions that the McArthur HYC deposit is at present Uneconomic.

Jan 1, 1984

By Haigh C. J

A process has been developed to recover at least 75% of the 2-naphthol contained in cobalt precipitate produced during the removal of cobalt from electrolytic zinc plant solution. The recovered material is in a form suitable for recycling to precipitate further cobalt from solution. The process consists essentially of two steps. The first involves leaching of cobalt precipitate in 400 g/l sodium hydroxide solution at 95¦C in the presence of an additive. The second step involves treating the filtrate from the leaching stage with sodium sulphide to precipitate cobalt as the sulphide thereby re- generating the sodium hydroxide for recycling. The process is only marginally profitable at present 2-naphthol prices, giving a net return of 20% on a capital investment of $1 200 000. A reaction mechanism is proposed which accounts for the 25% loss of the 2-naphthol in the recovery process.

Jan 1, 1977

By J Ralston, G B. N Kaggwa, L Huynh

Three polymers of varying chemistries and molecular weights have been used to depress talc in a model sulfide ore consisting of talc, pentlandite, and chalcopyrite. Flotation testing revealed a distinct hierarchy of effectiveness of the polymers in the depression of talc (HP Starch > Polymer-H > Dextrin WY) but a different order of effectiveness for depression selectivity (Dextrin WY > Polymer-H > HP Starch). Adsorption isotherm determinations for the gangue and valuable mineral phases indicate that the differences in selectivity are partially due to adsorbed amount variations, with Dextrin WY adsorbing marginally less on the sulfides than Polymer-H. Contact angle measurements on packed particle beds have revealed that the polymers have differing abilities to reduce the contact angle of talc and the sulfide minerals. An explanation of the observed trends is given in terms of the binding mechanisms of the polymers on the talc and the sulfide mineral phases. Order of addition effects between collector and depressant were also investigated using adsorption isotherm determinations for the polymers in the presence of a pre-adsorbed layer of collector and vice versa. Adsorption of the polymers was not affected by the presence of the collector. Similar results were obtained for the adsorption of collector in the presence of pre-adsorbed layers of polymer.

Jan 1, 2005

The underground coal mining industry in Australia has, in the last 30 years, undergone a revolution in production methods, techniques and results probably unparalleled elsewhere in the world. It can lay claim not only to leading the fight against pneumoconiosis and setting high safety standards generally, but has at the same time recorded comparatively high rates of increase in output per man shift. The latter has been achieved as a result of progressively matching the latest in production and underground transportation equipment with significant changes in seam development and extraction techniques and the introduction of stress-relieving extraction methods unique to the world scene. However, the seam recovery factor is still too low. Apart from the overall objectives of maximising production and commercial returns with the greatest safety, there must be a concentration towards maximising seam recovery, particularly in high seams. The search will continue for systems which are more completely continuous mining operations than the present stop-go operations, with longwall having long term potential in this regard, but as yet not sufficiently developed to engender confidence for its more general use. Techniques developed in the last 10 years make the development and extraction of deeper seams, perhaps to 2500 feet or more, feasible with continuous miners as well as longwall. The future will see a greater use of continuous shifts at mines, larger mines with more sophisticated management organisation, extensible belts at the face, better systems of immediate roof support at the face, remote operation of face equipment, greater use of roof bolts and the longer roof cables. With commercial confidence in coal mining now virtually assured to at least the turn of the century, the stage is set for heavy capital development in the industry which will undoubtedly include major research into the whole area of mining methods and equipment. We must ensure that our contribution to such research equitably matches other coal producing countries.

Jan 1, 1975