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By Welch B. J

Results obtained from linear sweep voltammetry at a graphite anode in cryolite-alumina melts are presented, together with those obtained from a chronopotentiometric investigation of the influence of dissolved metal. The results, consistent with both production information and other fundamental studies, showed that the following two reactions occur at the anode.Equation 1Al (dissolved) ? Al3+ + 3eThe latter accounted for some of the loss in currentefficiency.The rate of the former reaction is limited by the slow desorption of the CO from the electrode surface. Mass transfer conditions, and hence cell design, influence the importance of the second reaction.

Jan 1, 1976

By Viney K. G, Molinia A. B

The Hellyer Pb-Zn-Ag orebody in western Tasmania was discovered in 1983, and mine production at a rate of 250 000 tonnes/yr commenced in 1987. Expansion to 1 million/yr will be completed by early 1989. Exploratory development to gain access to the orebody, to provide a 20 000 tonne bulk sample for pilot plant work and to confirm mining conditions and orebody continuity was undertaken using conventional drill and blast techniques and large diesel trackless equipment. A ventilation shaft was raisebored. Development in this phase totalled 2 815 m and provided a sound basis for rapid expansion of the mine.

Jan 1, 1988

Under South Africanncoalmining conditions, in the majority of instances coal is won using the bord and pillar system without pillar extraction. Examples demonstrate that strata control policy has an effect on practically all aspects of mining: (1) the choice of mining method; ((2) degree of coal recovery; (3) effect on productivity; (L}.) integration of'-support techniques with other operations. This is confirmed by use of modern research results and computer simulation. Although the examples are drawn from South African experience, the conclusion is generally valid. Strata control problems cannot be solved in isolation but as an. integral part of overall mining policy.

Jan 1, 1971

By Wencai X

The conversion relationship between the Bond Work Index W1, specific output of mill and the energy efficiency of grinding were developed using dimensional analysis. These relationships were validated and the value of the conversion coefficient ~ was also obtained using a new installed device for measuring grindability of ores. With the new simulating algorithm the mill specifications, output and size distribution of grinding products of industrial mills can be calculated based on laboratory measured results. This algorithm is more accurate and simpler than any of the other empirical relationships. KEYWORDS: Conversion coefficient method, grindability, grinding Constant.

Jan 1, 1993

As previously advised, two of the raw materials for Wilsons Portland Cement are cement rock and limestone, both of which are quarried by the company. We now wish to give you a brief outline of these operations. The major material, cement rock, is obtained for our Portland Quarry at Tikorangi Hill, which is one mile west of the Works and a familiar landmark to travellers on state Highway 1. The geological structure of this deposit is involved, having intensive and irregular earth movements interpreted as large scale slumps where older rocks overlay younger rocks - numerous folds, faults and overthrusts are clearly visible.

Jan 1, 1978

By Pollack B

Four different major international copper projects are discussed. Three have recently come on line, and one is now being constructed. Each of the projects has encountered unique problems during the course of its development because of the legal requirements of the host country, together with its political and social needs. Nevertheless, there have been similar problems to solve. There were six common factors. 1. Quality engineering to be performed under the direction of an experienced project team. 2. Good communications. 3. Logistics. 4. Customs requirements. 5. Availability of capable construction labour. 6. Local material source capability.

Jan 1, 1983

By T Christopher

The Campamento Au/Ag deposit, discovered by Linear Gold in 2004, is located in Chiapas State, southern Mexico. The deposit is hosted by a sequence of Pleistocene-Pliocene andesitic volcanic rocks that have been intruded by a feldspar-phyric porphyry complex. Measured and indicated resources are 852 000 oz Au and 3.46 million ounces Ag with inferred resources of 374 000 ounces Au and 920 000 oz Ag. Resources were estimated at a 1.0 g/t Au equivalent cut-off. This is an ABSTRACT only. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By Collins BJ

Basic principles of stack injection are outlined and supported by calculations of thermal, chemical and aerodynamic requirements of an ideal 4000 tpd furnace. Calculated results show that these requirements vary with conversion of FeO, and that best performance may be expected with XFeO = 1.0 for which calculated production is 6919 tpd and coke rate is only 211 kg/thm. Calculations are also presented for a cycled injection. Discussions show that stack injection is a good way of overcoming softening problems and meeting the demand for iron in a fluctuating demand situation when it exceeds the maximum capacity of a furnace.

Jan 1, 1975

The Olympic Dam copper/uranium deposit was discovered in 1975 and subsequently commenced production in 1988 at a rate of 1.5 Mtpa using the sublevel open stoping mining method. In the following five years the operation underwent two optimisation programs that have increased mine output to the current rate of 3.4 Mtpa of ore. The pre-expansion requirements were annual development of approximately 10 000 m and the placement of around 500 000 m 3 /yr of cemented backfill. Previous to 1995 all hoisting was carried out by the Whenan shaft. The Robinson shaft was commissioned during that year to hoist increasing development mullock tonnages and to allow for increased ore hoist via the Whenan system. In 1996 WMC committed to expanding the output of the operation from 85 000 tpa (1996/97 budget) copper metal to 200 000tpa. The corresponding increase in mine ore production requires expansion to a nominal rate of 9.0 Mtpa. To achieve these increased tonnages, major hardware installations will include sinking of a third shaft capable of hoisting in excess of 9.5 Mtpa, a rail haulage horizon utilising automatic trains, expansion of the existing backfill plant to a production capacity of over 1.8 million m 3 /yr, increasing long term development rates to 15 000 m/yr and increasing the production drilling and blasting rates in line with the higher stope production output. This paper provides detail into the intended hardware (shafts, equipment, etc) of the mining system, and outlines the intended long-term operation and management of the operation. Of particular note is the background requirement of ensuring existing production targets are met to provide critical cashflows into the project during the period of expansion.

Jan 1, 1998

By Millett R

In January 1989, Western Mining Corporation (WMC) acquired the Perseverance and Rocky's Reward orebodies and associated infrastructure from the Agnew Mining Company (AMC). AMC had closed the operation in August 1986 and it had been on care and maintenance since that time. To return the operation to full production as quickly as possible it was planned to mine the Rocky's Reward orebody by opencut mining and treat this ore whilst the underground Perseverance workings were rehabilitated. Removal of overburden from Rocky's Reward thus commenced immediately with a target of April 1989 for delivery of the first ore to the concentrator. Extensive rehabilitation and upgrading of the concentrator commenced in late January, and continued in stages for an 18 month period. The first stage consisted of overhauling all existing equipment and completing a crusher upgrade commenced by AMC prior to the mine closure. During this period, it was necessary to recruit a workforce and re-establish all the required support structure. In late April the plant commenced commissioning trials on mullock. The first ore was crushed on the 11th May 1989 and concentrate produced the following day. From May 1989 to June 1990 the plant was producing nickel concentrate whilst further upgrading continued. During this period, the flotation capacity was tripled by installation of a bank of eight Wemco 144L flotation cells in rougher-scavenger duties and two 12 metre by 2 metre diameter columns in cleaning duties. The existing vacuum drum filters were replaced by two Larox PF32 pressure filters, eliminating the need for solar drying and allowing the flash drier to be fed direct from the filters. These, along with a major borefield expansion and further work to the crusher increased the nominal plant capacity from 600,000 to 1,000,000 t/yr. The plant achieved the 1,000,000 t/yr rate in April 1990.

Jan 1, 1991

By Hedge AS

Mt Whaleback is one of the largest open-cut iron ore mines in the world. Haulage cost is the single largest component of operating costs at the mine. Past studies towards alternatives to truck haulage have recognised the potential viability of an in-pit crushing and conveying system, dependent on the proportion of uphill haulage. After seventeen years of mining Mt Whaleback still contains more than 1,000 million tonnes of ore, needing 1,800 million tonnes of waste removal to extract it. What was once a quarry has now turned into a large open pit. The increasing proportion of uphill haulage, the need to replace ageing 120 ton and 200 ton trucks and the necessity of achieving and maintaining a higher stripping ratio prompted, in 1984, the initiation of a new feasibility study into the viability of alternative material handling systems. A semi-mobile in-pit crushing and conveying system was found to be most suited to Mt Whaleback conditions. The East Pit was selected as the best location based on the analysis of haulage profiles measured on long term incremental mine plans. A capacity of 22 million tonnes per year was considered achievable by the proven technology currently available. A detailed estimation of capital and operating costs of two systems - one with trucks only and another with trucks, crusher and conveyor - for the removal of 22 million tonnes per year of waste rock, showed substantial cost savings achievable by the proposed system. An economic evaluation of the two systems established the advantages of constructing the in-pit crushing and conveying system. This is now under construction and will be commissioned in early 1987. Experience gained in constructing and operating this installation is expected to influence decisions on similar installations in future.

Jan 1, 1986

Coal losses and dilution are among a coal mine operators everyday problems. These losses and unwelcome additions are generated primar- ily by the stripping operation and secondly by the coal mining opera- tion. Losses in yield due to preparation plant efficiency are a separate issue although they may be related to the coal mining operation. The losses and dilution can be traced to the following operations: (A) Overburden or Interburden Removal (1) Removal of coal edge by blasting operations. (2) Removal of top of seam by drilling/blasting. (3) Removal of top of seam by stripping equipment. (4) Removal of coal edges by stripping equipment. (5) Dilution due to stripping equipment. (B) Coal Mining (for the case where seams are ripped rather than shot) (1) Coal rib losses. (2) Coal rib dilution. (3) Coal left in floor. (4) Coal left in fault areas. (5) Floor material picked up whilst loading. (6) Spoil material picked up. (7) Spoil material from poor surface clean up. (8) Top of coal losses from final clean up. (9) Dilution from ripping too deep, i.e. into floor below seam. (10) Dilution from interseam parting. The losses apply to both the single seam and multi seam oper- ations although they are more pronounced in the multi seam mines than in a single seam mine due to the position of the coal which is lost. For example in the multi seam case the top seam coal edge loss is final. There is little chance of recovery with the subsequent pass even if the extra rehandle is done. However conversely the lower seam is relatively easy to locate as it has nowhere to go being contained by the lowwall bench.

Jan 1, 1988

By Orr T. O H, Hodkinson I. P

The Mount Leyshon breccia complex is located on an east-west trending contact between Cambrian metasediments (Puddler Creek Formation) and the Ordovician Mt. Leyshon granite, within the Lolworth-Ravenswood block. Detailed mapping of basement, breccia and igneous unit geometries, crosscutting relationships and resolution of complex chronological relation- ships indicates that the Mt. Leyshon breccia complex should be reinterpreted as a number of intrusive igneous and breccia phases. This is in contrast to a previously published account of a largely extrusive geological setting. A complex structural history dominated by strike-slip faulting is preserved within the basement units and is responsible for controlling the location and geometry of the breccia complex. The breccia complex has been divided into brecciated "basement", "main pipe breccia" and an "interactive magma/breccia sequence". Main pipe breccia (the most voluminous unit within the complex) has been re-interpreted as an intrusive breccia, rather than an extrusive or vent fill as previously proposed. This is based on the following evidence: 1. the facies characteristics of the main pipe breccia, 2. the presence of a progression from cohesive basement through to mixed breccia, 3. minor basement block rotation during breccia complex formation, and 4. the partial structural control on the main pipe breccia. No unambiguous evidence exists for pyroclastic facies, although the fine breccia variant commonly contains "accretionary structures", very similar in appearance to those formed during pyroclastic processes. In view of 1. above, they are interpreted to be the product of particle accretion during upward movement of fragmental material, although an origin due to altera- tion cannot be completely discounted. heir occurrence in the main pipe breccia indicates that their presence within a rock unit is not necessarily diagnostic of nvroclastic nroeesses.

Jan 1, 1991

Since the mid-1970s there has been increasing attention in the South Pacific area to the problems of disaster planning and management at various levels, mainly from regional through national to provincial levels. This concern has been expressed in a variety of ways. International aid donors have made their views known in commentaries or reports. Seminar discussions have been organised at the regional level (such as at Suva, 1976; Suva, 1979; Tonga, 1982 and Suva, 1983). In-country training workshops or meetings have been organised (such as the "Disaster Preparedness Workshops for Non-governmental Organisations" in Pacific countries, jointly run by the Australian Red Cross and the Australian Overseas Disaster Response Organisation and supported by AIDAB, in Vanuatu, 1985; Solomon Islands, 1985; Western Samoa, 1986; Fiji, 1987; Tonga, 1987; Lae and Port Moresby, Papua New Guinea, 1988) or held outside the area (ACDC at Macedon, 1981). Training and/or research institutions have been established such as the East-West Center, Honolulu, and the Asian Disaster Preparedness Center, Bangkok (more Asia oriented). As part of the East-West Center Pacific Islands Development Program's Disaster Preparedness Project, a detailed manual of disaster preparedness and response procedures and guidelines has been published (Carter, 1984).Table 1 and figure 1 show that the area with which this discussion deals is a highly fragmented island region, comprising many small, some minute, nations, both in land area and population. The social structure is predominantly rural, with the capital city often making up a small part of the total population and, in absolute terms, little more than a small town. Outside the administrative and governmental activities concentrated in the capitals, in most cases the indigenous inhabitants depend on subsistence agriculture or on usually a monocultural concentration on a cash crop, most frequently coconuts.

Jan 1, 1990

By P Mackenzie, S Askraba

Pyritic black shales present several challenges to mining operations due to their reactivity. The risk of predetonation of explosives in blastholes, spontaneous combustion of fractured rock and the generation of acid rock drainage must be managed for safety, operational and environmental reasons. At the Mt Tom Price iron ore mine there is a need to differentiate between reactive and non-reactive shale in order to assess risk and make decisions on materials handling. The current approach uses a combination of stratigraphic information and sulfur analysis of blasthole samples. The time taken for sulfur analyses can be a constraining factor on mine production. A new instrument developed by ANSTO Minerals overcomes some of the current limitations on material characterisation. The IOR MeterÖ measures the intrinsic oxidation rate (IOR) of samples of broken rock by circulating air through a sample and monitoring the oxygen concentration over time. The intrinsic oxidation rate is given by the rate at which oxygen is consumed per unit mass of sample, in units of kg(oxygen) kg-1(sample) s-1. Up to eight samples can be measured simultaneously and the turn-around time can be less than 12 hours. Initial trials of the IOR MeterÖ were undertaken at the Mt Tom Price mine in July 2004. The measured IOR of samples from a blasthole pattern spanning different geological layers containing black shale, correlated strongly with the expected reactivity of those layers. The correlation between position in the geological sequence (detailed stratigraphy) and IOR MeterÖ results were much stronger than that between stratigraphy and sulfur content. The new characterisation technique offers the prospect of improving the ability to identify and predict the behaviour of reactive black shales and to differentiate more precisely between different materials. Mine planning, operations, safety and environment all stand to benefit from the success of further trials of the technique.

Jan 1, 2005

By V Salamakha, K Tang, G Si, J Oh, X Wu, Z Yu

The development of machine learning (ML) algorithms has led to promising advances in lithology classification from geophysical logs, which is indispensable in various underground engineering applications. Due to the resolution and availability of logging data, ML models can recognise lithology change in a definitive form at a fine scale (0.01 m). While the results from previous models indicated promising performance in the classification, the robust capability of generalising lithology prediction to unseen data may not be true for those models due to: (1) data bias caused by borehole selections, and (2) logging data mismatches among boreholes. To this end, this paper aims to investigate the generalisability of ML models and seek potential improvements. Four ML models were selected to be trained on single reference boreholes to test the other (as unseen) in the same region, which includes Support Vector Classifier (SVC), Random Forest (RF), eXtreme Gradient Boosting (XGBoost) and Residual Neural Network (ResNet10). The data set involves 11 boreholes from a coalmine in north-western Bowen Basin (Queensland, Australia): density, gamma ray, neutron, and sonic logs are selected as inputs. Additionally, a data adaptation method is applied for better generalisability. The results show that there is an accuracy trade-off between the same borehole and unseen boreholes across the models. It also indicates a 16–43 per cent reduction in model performance when generalising the predictions (in marco-F1 score), while the adapted data sets can contribute to a around 12 per cent improvement. This study provides a fundamental understanding of the model generalisability when using a single borehole, essential for further correlating boreholes for regional lithology classification; the adaptation method can improve the generalised accuracy, and reduces the labour to label lithology, which facilitates the identification of gas storage mechanisms, geological/geophysical modelling, and stratigraphic analysis.

Sep 1, 2024

By P K. Seccombe

Sulfur isotope data are available for 18 low-sulfide orogenic gold systems hosted by the Lachlan Orogen in eastern New South Wales including those in the northern Hill End Trough (Hargraves, Hill End, Napoleon Reefs, Sofala-Wattle Flat, Stuart Town, Windeyer) and adjacent areas (Bodangora, Lucknow, Springfield), the Gilmore Fault Zone (Adelong, Sebastopol) and those in the Parkes district (Calarie, London-Victoria, Wyoming One-Myall United). Other systems for which data are available include Mongarlowe, Murrumbateman, McMahons Reef and Yalgogrin. The d34S data for many of the larger deposits including Adelong, Bodangora, Calarie, Hargraves, Hill End London-Victoria, Lucknow, Sebastopol, Sofala-Wattle Flat, Stuart Town and Windeyer (Group 1) are all very similar and have average d34S values close to zero per mil. Near 0ë d34S signatures imply that sulfur was sourced from a magmatic reservoir. Furthermore, the large Wyoming One deposit has d34S values significantly below 0ë with the data implying that sulfur was sourced from a fractionated magmatic reservoir. Group 2 deposits include those at Mongarlowe, Murrumbateman, Napoleon Reefs, McMahons Reef and Springfield . These deposits have average d34S values significantly above 0ë. With the exception of Springfield where sulfur was probably sourced from the Hill End Trough, Group 2 deposits are an order of magnitude smaller than Group 1 deposits. The results suggest that many of the larger orogenic-related gold deposits hosted by the Lachlan Orogen in eastern New South Wales derived sulfur and gold as a result of leaching and recycling from a magmatic reservoir. Alternatively, there may have been direct mantle-derived input; however, this is considered unlikely on geological grounds. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

Naturally occurring rutile contains approximately 96 per cent titanium dioxide (Ti02), whereas the product from the ilmenite reduction - aeration upgrading process, "sub- rutile", contains only about 92 per cent Ti02. Therefore, there is an-incentive to lift the quality of sub-rutile by removing transition metal impurities, in order to supply a more valuable and marketable product, particularly for the chloride-route pigment producers. Previous studies have shown that the sub- rutile residual iron content (of approximately 4 per cent) and residual manganese content (of approximately 1.5 per cent) are chemically bound within the titanium-oxygen mineral lattice, and are most difficult to remove by simple leaching methods. In attempts to improve their eventual removal, a simple process modification has been successfully investigated in laboratory scale trials. Transition metal sulphides have been formed in the reduced ilmenite by using a "sulphur-doping" technique during reduction (metallization) conducted at controlled temperatures. These compounds give rise to greatly enhanced dissolution of the unwanted transition metal constituents during the sub- sequent aeration and acid washing steps. Thus, the manganese content of the final sulphur- doped, reduced-aerated, acid washed product can be reduced by about 90 per cent to 0.2 per cent, and the iron content by about 70 per cent to less than 1.5 per cent, in comparison with those levels remaining after unmodified pro- cessing. Additions of sulphur as low as 2 per cent of the weight of the ilmenite have proven satisfactory. Preliminary studies reveal that the modified process is economically feasible. These results suggest that the sulphur content of the Collie coal char used in an unmodified reduction step must exert some influence on the eventual removal of iron and manganese during aeration and acid washing.

Jan 1, 1973

IN preparing a sample of refined lead for analysis, it is necessary to exercise great care to avoid contaminatien, as the amounts of the various impurities present in market lead are very small. The lead is cut into convenient-sized pieces, which are flattened out with a hammer, so that, on putting tbe lead through the rolls, strips of about one inch in width and Ij32iu. in thickness are obtained. The hammer, anvil, and rolls should be carefully cleaned before use. The strips oflead are cut up into pieces of about one inch square. If the lead is roIled out very thin and cut up into small pieces there is considerable risk of the acid acting too violently and causing loss through frothing over. A blank assay on all the reagents used should becarried right through, along with the actual analysis of the lead. This is of the greatest importance, particularly with regard to the determination of such impurities as iron, sulphur, arsenic, and antimony.METHOD OF ANALYSISThe method of analysis is as follows: Take two lots of 600 gIm. each, place in a large beaker, add 2000 cc. distilled water and 750 cc. concentrated HNOs. Cover with a clock glass and warm on a hot asbestos plate until all action ceases. (If any white precipitate forms, filter off and examine for Sn and Sb.) Add 420 cc. H2S04 (1-1), and stir well while adding the acid: Allow PbS04 to settle, pour off the clear liquid into a 12-in. porcelain evaporating dish, and start evaporating down, the dish being placed directly on an 8-in. ring gas burner. Wash the precipitate three times by decantation using approximately 1200 cc. hot water for each washing, and stir well...

Jan 1, 1922

By Mercer J. K

Shaft guide reaetions in No. 2 Ore Shaft, Mount Isa Mines Limited, have been reduced to a low order of magnitude by the use of self aligning slippers and spring loaded guide rollers. Before commissioning the new 4500 h.p. winder at No. 2 Ore Shaft in January, 1958, it was realised that the increase in hoisting speed from 1800 to 2650 ft/min. would cause high shaft guide reactions which would be increased further when the elongated 12 ton skips replaced the existing 7 ton skips.The Cambridge accelerometer was used to measure the shaft guide reactions and to confirm that progressive skip modifications were resulting in reduced reactions. The skip modifications were completed in 4 weeks.INTRODUCTIONIn 1956 an expansion programme was initiated to increase production from 5000 to 11000 tons per mining day. At No. 2 Ore Shaft the 1500 h.p. winder operated at 1800 ft./min. and hoisted a 7 ton capacity skip. The new 4500 h.p. winder operates at 2650 ft./min. and hoists a 12 ton capacity skip. No. 2 Ore shaft is equipped with two 5 ft. by 5 ft. hoisting compartments (Fig. 1) and any increase in skip capacity meant an abnormally long skip. This feature, coupled with the increase in hoisting speed, could have caused very high shaft guide reactions. Before commissioning the 4500 h.p. winder, a Cambridge accelerometer was borrowed from the University of Adelaide and was used to measure the shaft guide reactions. Rather than carry out an extensive shaft timbering programme, the skips were modified by fitting cushioned guide shoes and spring loaded guide rollers (Fig. 1). Testing commenced in February, 1958, and was completed in March, 1958.

Jan 1, 1961