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The purpose of this paper is to describe briefly the clean-up procedure adopted at the treatment plant of Paringa Mining & Exploration Co. Ltd. at Kalgoorlie, Western Australia.A clean-up is carried out at the end of each four-weekly period, and all figures subsequently quoted are taken from a summary of the results of the eight periods ending 8th April, 1941.STRAKING AND AMALGAMATIONOf the total gold won 22· 77 per cent. is recovered by straking and amalgamation.The discharges of both the primary and secondary ball mills are pumped over a common set of five ten feet long by three feet wide straking tables, set at an angle of 10 degrees, before passing to the 14 ft. diameter by 30 ft. by 8 ft. Dorr bowl classifier.This classifier is in closed circuit with both of the ball mills, hence the whole of the circulating load of the fine grinding circuit, which approximates to 800 per cent., is required to pass over the straking tables. For convenience in handling, corduroy cloths are used in lengths of 5 feet, and are picked up and replaced by clean cloths at four-hourly intervals. The avaHable straking area provided is, therefore, 900 square feet, or approximately 3·30 square feet per original ton per 24 hours. The density of the pulp passing over the tables is about 54 per cent. solids, the area provided per ton of pulp per 24 hours being approximately o· 2() square feet. Washing of the corduroy cloths is carried out on the day shift only, about 800 lb. of strake concentrate being obtained daily.This concentrate is sluiced from the washing trough on the elevated straking floor to the 4 ft. by 2 ft. diameter amalgamating barrel located in the gold room below.The barrel is of standard design and is rotated at 25 r.p.m. It is charged each day at 4 p.m., and grinding of the strake concentrate at approximately 70 per cent. solids is continued until 8 a.m. the following day. The barrel is then stopped, and the contents generously diluted with water, and after a few minutes rotation some of the slime is carefully poured off through the small discharge hole. This dilution and decantation is repeated until a good deal of the valueless slime has been removed.A further quantity of strake concentrate is then charged to the barrel, which is allowed a further 16 hours grinding period. After again diluting and decanting, approximately 25 lb. of mercury and 15 lb. of lime are added to the charge, which is then given an amalgamation period of 2 hours before cleaning up.By the process of grinding, diluting and removal of slime by decanting...

Jan 1, 1941

Small scale' table concentration tests have many critics. Many metallurgists consider that such tests are of problematical value because of the difficulties involved in conducting and interpreting them.Many kinds of small-scale ore dressing tests are difficult to conduct, and there is, perhaps, good reason for thinking that table concentration tests are amongst the most difficuIt.Interpretation of results from small-scale tests is the responsibility of the metallurgists and engineers in charge, and it is often held that small-scale table concentration tests are particularly difficmlt to interpret.TESTING FOR TABLE CONCENTRATIONOn analysis it appears that the difficulties encountered in conducting small-scale table concentration tests are of three fairly distinct kinds. Firstly, there are difficulties due inherently to the smallscale nature of the operations; for example the smaller width of all mineral bands on the table and the less complete separation due to the shorter length of travel between the feed and discharge points.Secondly, there are the effects of batch operation owing to the fact that the mineral particles behave differently during the initial period when the sample is just beginning to spread over the table, the middle period when feed and discharge are even and continuous, and the final stage, when the last of the sample has been added and the table is beginning to empty itself.

Jan 1, 1954

By R Edwards

Greater efficiency of capture of minus 6 mm to fine micron size particles of free gold is achieved with the Knelson Concentrator. The development, principles of operation and recent innovations relating to this enhanced gravity concentrating device are described by the inventor, followed by a description of application and consequent economic benefits in low grade alluvial gold deposits by a large scale user of Knelson Concentrators in the Northern Territory of Australia.

Jan 1, 1990

Operating well designed and maintained surface mine haul roads is the key to minimising truck haulage costs - in itself a significant component of the total cost per tonne mined. Existing and proposed mining operations are subject to scrutiny, both from the economic and environmental perspectives. In the long run, this scrutiny returns improved efficiencies and leaner, 'greener' operations. The focus of this evolving evaluation process should and will certainly fall on haulage operations - simply by virtue of their contribution to overall cost of operations and sustainability impacts. Whilst the end result - improved efficiency and contributions to resource sustainability - is not in itself problematic, it is the route, or process followed to achieve these improvements that needs to be carefully managed. We can be guided on this journey by our understanding of how a road is designed, and, critically, the technological solutions that can contribute to a company's sustainability initiatives. This paper examines the current state of mine road design and operation, together with the economic and environmental issues associated with under-performance. The paper serves as a basis for evaluating the economic and environmental contributions that technology in road design and management can deliver to sustainable operations. Utilising innovation is the key to long-term sustainability and as a means of positioning the industry for the imminent structural changes to the economy and environment, changes which are now within the operating life of many mines, haul roads and haul trucks.

Jun 1, 2010

By Moorrees C, Everson P

A detailed examination of carbonic acid leaching of crude iron-containing magnesia derived from Savage River magnesite shows that the rate and extent of magnesium and iron dissolution are affected by the slake time, leaching temperature, carbon dioxide pressure, pulp density and rate of agitation. In order to produce high-purity magnesia from the leachate by precipitation and calcination, the optimum leaching conditions, which are those that yield maximum magnesium and minimum iron dissolution, have to be a compromise in terms of the above parameters. For example, iron dissolution can be limited by leaching at an elevated temperature (-45°C), but because the maximum concentration of magnesium biearbonate decreases rapidly with an increase in temperature, it is necessary to use a low pulp density (-2 per cent solids) in order to achieve an acceptable degree of magnesium dissolution.Evidence is obtained that iron dissolution involves the formation of a soluble ferric bicarbonate complex that decomposes at elevated temperatures. The extent of iron dissolution is much higher than might be expected and it is probably because the original magnesite contains siderite in solid solution with the magnesite; on calcination, the hematite whieh is formed is in a very finely divided condition and hence more reactive.

Jan 1, 1983

Two outstanding features of the world's young orogenic belts are their very consistently developed suites of volcanic rocks and, it is now recognized, the rather distinctive types of mineralization they contain. Further,it is now known that these volcanic suites exhibit evolutionary patterns, and it is beginning to appear that this evolution in style of igneous activity is accompanied by a parallel evolution in style of mineralization. Specifically,podiform chromite and stratiform pyritic copper ores of Alpine ultramafic/submarine tholeiite association progressively give way to more zincian stratiform ores associated with andesites and dacites, "porphyry copper"deposits developed in quartz dioritic to dacitic plugs, and lead-zinc stratiform ores related to increasingly pyroclastic, rbyodacitic volcanism.This "coupling" is exhibited well in the volcanic arcs - the young orogenic zones-of the SW Pacific which, with tbeir abundance of youthful and commonly unmetamorphosed materials provide an opportunity for investigating, through a petrochemical approach, the possible significance of ores in the elucidation of mechanisms of lava formation. It is suggested as a working hypothesis that in the orogenic (voleanic arc) regime ores are associated with the cale-alkaline rocks because these cale-alkaline rocks are in part generated by exhalative loss, and the ores are among the manifestations of tbat loss.

Jan 1, 1978

By J E. Everett, K Jupp, T J. Howard

Precrusher stockpiles are designed principally as buffers to decouple the mining and processing operations. They are usually paddock dumped or dumped over a face to form fingers by dumping haul truckloads and reclaimed by front-end loader in an ad hoc manner. In addition, they are often not built and reclaimed to completion but continually added to and extracted from. While this design suits the mining process and is operationally simple, there can be little confidence in the grade of the ore that is fed to the crusher. This makes short-term grade control difficult and reconciliation back to the mine face imprecise. Well-designed and operated precrusher stockpiles can overcome these deficiencies.The different types of grade variability in mining are reviewed in this paper. The role of precrusher stockpiles in reducing short-term variability is discussed and their ineffectiveness in addressing long-term variability is highlighted. Their role in removing the serial correlation of the extracted ore is explained.Precrusher stockpiles carry out the four, at times competing, objectives of storing, buffering, blending and grade separation. Their actual design and operation result from a compromise between these four competing roles. These roles are explained in detail along with the advantages and disadvantages of the different types of precrusher stockpiles. The recommended blended-in blended-out stockpile (BIBO) design is discussed in detail. Matched pairs of BIBO stockpiles of limited tonnage built and reclaimed to completion retain the best knowledge of grade and facilitate the reconciliation process.Simulation studies have been carried out on BIBO stockpiles to understand the blending opportunities of the different options for building and reclaiming, and to identify the methods that achieve maximum blending. It was found that laying down rows in one direction and reclaiming across these rows gives the best blending. Matching the width of the reclaim face with the daily crushing requirements is also important. The length of the rows, and building in either one or two layers has little effect on the extent of blending.Appropriate design can provide a reasonable compromise between the four objectives.CITATION:Jupp, K, Howard, T J and Everett, J E, 2013. The role of precrusher stockpiling for grade control in mining, in Proceedings Iron Ore 2013 , pp 203-214 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

Recovery of Vanadium from Stone Coal Through an Environmentally Friendly Acid Leaching Technology

Sep 13, 2010

By R F. G Phelps, C J. S Sangster, G S. C

Narrow vein metalliferous mining in the UK has a long history, but there are only two small gold operations at present (Clogau and Gwynfynydd). Four further gold properties are at various stages of evaluation (Calliachar, Cononish, Curraghinalt and Maestryfer). The South Crofty tin mine closed earlier in the year. Narrow veins are exploited underground and are challenging to evaluate because of their geologically variable nature. They are generally less than 3 m wide, discordant, of variable dip (

Jan 1, 1998

By Prasad S, Chatterjee S. S, Singh B. N

The chief source of metallic iron for steelmaking in India, even today, as it is in most part of the world, is through blast furnace (BF) route. This will continue to be so in the foreseeable future as the technology is most efficient, both process and economy-wise. However, as the world is moving into the next millennium, a critical look is being given to the process primarily because of its non-ecofriendliness (as a result of extensive use of metallurgical coal/coke requiring elaborate pollution control facilities) and huge investment involved for supporting facilities like extensive raw material preparation and handling, agglomeration, adoption of sophisticated instrumentation and control system, etc.

Jan 1, 1999

By R S. Hamilton, S R. Grobler, M E. Greenway

The progressive development of underground mines frequently requires increased ventilation flow rates through mine shafts that were originally designed for lower rates. Where these shafts are equipped with conveyances running on rope guides, the higher air flow rates adversely affect the lateral motion of conveyances and could cause conveyances to collide with the shaft walls or with one another. Hence the safety impact of higher flow rates requires proper assessment. This can be undertaken by use of computer simulation models to predict the dynamic motion of the conveyances during hoisting. It can also be undertaken entirely by field testing, or using a combination of simulation and field testing. This paper discusses these techniques and makes cautionary remarks about setting an absolute generic maximum flow velocity for all shafts. Approaches to assess existing shafts for increased ventilation flow rates are also discussed.CITATION:Greenway, M E, Grobler, S R and Hamilton, R S, 2017. Safety impact of increased shaft ventilation on rope guided conveyances, in Proceedings Australian Mine Vent Conference 2017, pp 261–270 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 28, 2017

By S L. Birch, M D. Raven

X-ray diffraction (XRD) is a well-recognised, invaluable tool to provide detailed mineralogical analysis on a wide range of applications. However, despite the wide use of XRD in many industries, there are currently few recognised standard methods for quantitative XRD analysis and none for the analysis of iron ore materials. An initial step towards preparing a standard method for XRD analysis of iron ores was to assess the current level of precision achievable by practitioners of the technique by organising an international round robin. Six Australian chemically certified iron ore reference materials were distributed to over 100 international participants. A summary of over 120 data sets from more than 80 participants is presented. Results show that a bewildering array of techniques for sample preparation, data collection, phase identification and quantification were used. The reported values for the main components (ie hematite, goethite, quartz and amorphous content) show not only extensive variability between participants but also within participants’ results for those who submitted multiple data sets. The disappointing low level of precision exhibited for a relatively simple mineral phase system indicates a strong requirement for development of a standardised procedure.CITATION:Raven, M D and Birch, S L, 2017. Summary of results of an international X-ray diffraction round robin for mineralogical analysis of iron ores, in Proceedings Iron Ore 2017, pp 599–604 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

The fundamental features of the Cable Belt Conveyor are described. It consists of independent wire ropes which take the transmission load, and a special reinforced rubber belt which acts 'only as carrier of material. The first cable belt installation in Scotland is described briefly.A general description of this type of conveyor is followed by details of the rubber belt construction and the method of joining sections of belt. The drive unit, the tensioning arrangements and the power requirements are outlined.Applications of the cable belt conveyor are 'dealt with and its advantages summarized.INTRODUCTIONThe cable belt conveyor, as its name implies is a conveyor composed of steel wire ropes and a rubber belt, in which the ropes serve as transmission members, and the belt merely as a carrier of materials.Over the years it has been the ambition of engineers connected with the belt conveyor industry to develop a conveyor that would use the high tensile strength of wire ropes to transmit the stresses involved in the driving of large capacity belt conveyors. Different lines of approach have been adopted in the United States of America and in Britain. In the United States the method adopted was to introduce a large number of flexible steel cables into the rubber belt itself, in the form of a closely packed layer arranged across the full width of the belt. These small cables are packed in a nylon filled duck envelope which is incorporated in the rubber belt. This is commonly known as the steel corded conveyor belt. In Britain, a Scottish firm developed a rubber belt conveyor in which full advantage was taken of the high tensile features of wire rope, without embodying or moulding the wire rope into the rubber belt, so that the two wire ropes and the rubber...

Jan 1, 1957

By P E. H Gregg, R B. Stewart

Restoration trials established in 1985 included a treatment where pasture seed was sown directly into fertilised tailings. After seven years, accumulated soil data was used to compare physical and chemical changes in the plots with the original tailings control area nearby.   A known amount of annual fertiliser additions over the seven year period allowed nutrient balances to be calculated. Losses of S and P were relatively high at 72% of applied S and 28% of applied P. Calculations show that at least 240 kg N/ha/yr was fixed by clovers and this is likely to be under-estimated given that the calculation did not allow for leaching losses.

Jan 1, 1993

By J S-L Chan, J C. T Chu

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. In its industrial days, Hong Kong’s mining industry was thriving, benefitting from the mineral occurrences that are related to the regional Mesozoic igneous and hydrothermal activities.Hong Kong was unique in the mining sector as it was so small in landmass but with such a rich variety of commodities discovered. There are records of commercial exploitation of tungsten, iron, lead, silver, beryl, graphite, quartz, feldspar and kaolin in Hong Kong before 1981. Of these, most of them are located in the New Territories.The Lin Ma Hang lead mine was operated in the 19th century by the Portuguese and was abandoned in 1958. The Needle Hill tungsten deposit was discovered in 1935 during construction of the Jubilee Reservoir. The Ma On Shan iron mine, which operated for 70 years between 1906 and 1976, was the largest mine in the history of Hong Kong. The Mui Wo lead-silver mine on Lantau Island can be traced back to the Ming Dynasty 400 years ago. West Brother Island, which extracted graphite from 1952 to 1973, was levelled in the 1990s to make way for the development of the airport at Chek Lap Kok.Nowadays, these mines are disused and abandoned, but this extended abstract reviews the legacy of the mining industry in Hong Kong to understand the impact that it had on techniques, the economy and culture.CITATION:Chu, J C T and Chan, J S-L, 2015. The history of mining in Hong Kong , in Proceedings PACRIM 2015 Congress, pp 109–114 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

The final criterion of the payability of gold production must necessarily be the cost per ounce. While the cost per ton or per cubic yard is a convenient comparative measure, and a necessary one for operating control, it is obviously insufficient as a means of judging whether or not, in a given set of circumstances, a gold production operation can be carried out economically.. An alternative method, which brings in the ton or cubic yard as a unit of measurement, is to determine what the cost per ton or per cubic yard is equivalent to, in terms of pennyweights or grains at the ruling price of gold. But this is really only another way of stating that the cost per unit of product is the ultimate measure of profit or loss, and not the cost per unit of input material.It is accordingly of interest to determine the cost per ounce of Australian gold production and to compare it with the cost of gold production in some other parts of the world which are important gold producers. The attached Appendix, which it, is believed comprises practically all Australian mines of any consequence (except Big Bell and the de Bernales group, for which. insufficient information is available), shows this information for financial years ending between 31st December, 1938, and 31st December, 1939, the production of gold dealt with being some 1,084,000 oz., or 66% of the total Australian production for 1939. For comparative purposes the cost of producing gold at representative mines in New Guinea, New Zealand, Fiji, Canada, India, West Africa and Rhodesia, and the average cost of production at the Rand mines, has also been included.It is obviously necessary in a study of this sort that the basis of the cost should be clearly set out and should be as nearly as possible uniform. The "cost" as calculated herein comprises the following items:

Jan 1, 1940

By A Vigar

Siliceous float sampled in the Wafi River by CRA Exploration in 1977 assayed 22 g/t Au and 89 g/t Ag. Follow up in 1979 led to the discovery of outcropping mineralisation now known as Zone A. Soil sampling in 1980 and 1982 outlined a 350 m by 400 m area of >1 g/t Au at Zone A and smaller areas at Zones B, C and D. Thirty-one diamond holes were drilled between 1983 and 1986. Elders Resources continued evaluation of Zone A under a farming agreement between 1988 and 1990. Mineralised diorite porphyry was discovered in 1990 by drilling conceptual targets 500 m to the NE of Zone A. Evaluation of this copper-gold porphyry became the main focus of exploration for the next five years and a substantial resource was outlined within a single, discrete body. A review of the paths to project development in late-1995 highlighted the need for increased resources via a second mineralised porphyry and/or an additional resource of Zone A size to make the project viable. It was concluded that the greatest immediate potential for increasing the resource inventory was at Zone A itself where the mineralisation was open in all directions. Detailed mapping and drill core logging lead to the geologic and structural models being revised in early-1996. The objectives of the 1996 drill program were to extend the Zone A resource, to test for continuity of mineralisation westwards to Zone B, to better define the controls on gold mineralisation and to delineate high-grade blocks amenable to selective mining by infill drilling and 3D orebody modelling. The last drill hole of 1996 intersected 30 m at 4.6 g/t Au from 273 m in WR147, deep in the western limit of Zone A. This intersection gave impetus to continue drill testing towards Zone B. In April 1997, hole WR154 was drilled 100 m to the west of WR147 and intersected 20 m at 8.5 g/t Au and 70 m at 11.0 g/t Au at similar depths near the diatreme contact. Drill hole WR155 was then drilled 100 m to the west of WR154 and intersected 86m at 5.7 g/t Au and WR159 drilled 100 m to the south of WR154 intersected 36 m at 9.2 and 50 m at 4.4 g/t Au. A significant zone of essentially blind, high-grade gold was discovered beneath the dipping flank of a poorly mineralised diatreme in what became known as the Link Zone. The long history of evaluation has been driven by the conviction of a dedicated team that a mineralised system of the size and complexity of Wafi with strongly anomalous gold in outcrops over the entire Mount Golpu area must contain economic mineralisation hidden within it. The keys which redefined the direction of exploration and lead to the discovery of the blind, high-grade gold Link Zone were, in our view, the critical review of the project status, detailed geological and structural studies to define clear exploration targets and the necessary funding and persistence to carry this through.

Jan 1, 1999

The polymerisation of silica is found to interfere with thickening of leach pulps at Risdon. Hydroxyl ion and fluoride ion are catalysts for this reaction. Stability may be obtained by the presence of aluminium or ferric iron.INTRODUCTIONAt Risdon it has been a matter for satisfaction that, although the concentration, of silica in leach solutions is appreciable, it has ,given relatively little cause for anxiety.However the margin of safety is small and some intermittent difficulty in thickening leach residues during the last few years has prompted a further investigation, certain parts of which are described here since they may be of general interest.Leaching and ClassificationThe method of leaching which is practised at Risdon has been described in, two previous articles, by W. C. Snow (1936) and S. W. Ross (1949) to whiCh the reader is referred for detail. Continuous leaching is practised, the circuit solution flowing through five air agitated pachucas in series, to tlie first of which calcine is added. The circuit solution, which also enters the first pachuca, consists of spent electrolyte from, the electrolytic cells and contains 90 -gm. of free sulphuric acid per litre. About 80 per cent of the solublezinc oxide is extracted from the calcine in the first pachuca, at a pH value of 1-1 (7 gm.H2S04/1 at 55 to 60°C. In the later pachucas the leaching rate decreases sharply. The leach pulp which leaves the 'last pachuca contains an average of 17 gm. of free acid per litre, corresponding to a pH of 1.7; but, owing to variations in extraction, the end point fluctuates between 2.5 and 0·3 gm. free acid per litre (pH:> 1·5 to 2.5). The average flow through the plant is 1,250,000 gallons per day, which corresponds to a production of 300...

Jan 1, 1958

Peculiarities of Quartz Distribution in Zinc Flotation Products and their Influence to Mineral Processing

Sep 13, 2010

By R Bezuidenhout, J K. Claflin, S R. La Brooy, G M. Wardell-Johnson, A M. Claflin

Best practice gold (Au) recovery is not an accident. It is the result of proper operation of the adsorption circuit using good quality activated carbon. A conventional carbon ‘regeneration’ kiln cannot process carbon fouled with floatation chemicals and achieve best practice gold recovery. A high temperature, controlled atmosphere carbon ‘reactivation’ kiln is required. Carbon circuit design allows for low quality carbon by increasing the carbon advance rate, carbon inventory and setting an easily achieved and commercially prudent solution tails target of typically 0.02 ppm Au. These design criteria are important to ensure the commercial success of the gold operation in the first instance. Most carbon ‘regeneration’ kilns will produce carbon of sufficient quality to achieve a 0.02 ppm Au solution tail in the absence of flotation reagents.Best practice recovery is considered less than 0.005 ppm Au in solution reporting to tails. If the carbon is heavily fouled, as it will be if used for flotation leach circuits, then a conventional carbon ‘regeneration’ kiln cannot achieve world’s best practice recovery no matter how much carbon is advanced. World’s best practice gold recovery requires a carbon ‘reactivation’ kiln and the whole of the carbon circuit operated optimally.The competing reactions occurring in a carbon ‘reactivation’ kiln are presented and discussed in terms of kiln design and operation. New industrial and laboratory findings are presented to fill gaps in the body of knowledge and identify that a large reduction in residence time at temperature will not significantly impair reactivation, but can significantly reduce the size and cost of the kiln or permit significantly increased carbon throughput.The carbon circuit performance is modelled in terms of gold adsorption, as a function of carbon reactivation using ‘simulation of carbon-in-leach/carbon-in-pulp circuits’ (SIMCIL). An optimum carbon advance rate is identified as a function of optimum carbon ‘reactivation’ kiln operation to achieve 0.005 ppm Au in the tails solution. A conventional carbon ‘regeneration’ kiln is compared to a state-of-the-art carbon ‘reactivation’ kiln in terms of reactivation performance and gold recovery.CITATION:Claflin, J K, La Brooy, S R, Wardell-Johnson, G M, Bezuidenhout, R and Claflin, A M, 2017. Optimising carbon reactivation kiln operation, in Proceedings MetPlant 2017, pp 346–364 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 11, 2017