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By Francis Foley

INTRODUCTION DURING the summer of 1919, the late Dr. Henry M. Howe, then Chairman of the Division of Engineering of the National Research Council, organized a committee to obtain a better insight into the behavior of carbon steels under the influence of various treatments by a proper control of the time and temperature factors. During part of 1919 and 1920, work in connection with this investigation was carried on at Dr: Howe's laboratory at Bedford Hills, N. Y. During 1921 there was an enforced hiatus in the work, occasioned by the illness which resulted in Dr. Howe's death in 1922. At that time one of the present authors compiled the results of the work which had been done and this was published in 1923.1 Following the publication of this work, it seemed desirable to carry on in the same manner an investigation of steel of C 0.75 per cent., which steel had not been as fully treated as had the other steels of C 0.34 and 0.52 per cent., used in the published work. Furthermore, certain unusual results have been obtained in the treatment of C 0.52 per cent. steel by quenching and drawing, and repetition of this part of the work seemed advisable. This has been done during the past year at the School of Mines and Metallurgy of the University of Missouri at Rolla, Mo., by a cooperative undertaking of the School and the Bureau of Mines, working in conjunction with the Committee on Heat Treatment of Carbon Steels of the National Research Council. The present paper is a report of the results obtained in this work.

Jan 2, 1926

By Thomas E. Marshall, Michael Dunn, L. J. Prosser, Anthony T. Iannacchione

During the past few years, the Pittsburgh Research Laboratory of the National Institute for Occupational Safety and Health (NIOSH) examined and characterized conditions at a majority of the underground stone mines in the United States. Observations at these mines revealed a limited degree of roof monitoring beyond visual inspection and sounding. When monitors are used, they typically require the miner to measure movement at the roof. If conditions are unstable, the mine rmay be in a hazardous situation while recording data. Based on this scenario, researchers surmised that a simple, inexpensive monitor with the capability of recording data at a distance from the mine roof would be a safer way to gain the information. Additionally, more widespread use of monitors could potentially lead to better understanding of roof movement in general. A monitor to meet this need was designed, tested, and subsequently improved as experience was gained in its’ use at a number of underground stone mines. The Roof Monitoring Safety System (RMSS) can provide an initial indication of movement in roof beams. By understanding and measuring roof movement in underground mines, the potential for injuries and fatalities to mine workers from falls of ground can be reduced. Also, officials at a mine with a history of data are better prepared to make a decision on remedial actions in the event of ground falls. This paper will outline the evolution of the RMSS and how it can be used in a comprehensive proactive ground control safety program. Also included is a case history describing how the RMSS was used in an evaluation of the effectiveness of a mechanical impact scaling machine at an operating limestone mine.

By S. Lolon

Spontaneous combustion in underground coal mines has become a serious problem particularly in the caved area (gob). Recent statistic has shown that approximately 17% of total 87 underground coal mine fires in the United States are attributed to spontaneous combustion (De Rosa, 2004). Spontaneous combustion results from a self-heating process in exothermic conditions. The accumulated heat, if not removed, is conducive to the rapid increase of temperature and may result in mine fires or explosions. It is well accepted that the interaction between oxygen and coal substances is the main cause for spontaneous combustion, while other factors such as pyrite, moisture, and bacteria play a secondary role to the self-heating of coal. Therefore, only coal oxidation is considered in this study. Coal oxidation occurs as coal comes into contact with air. This process involves complex phenomena in terms of heat transfer, chemical surface absorption, and energy balance related to inherent properties of coal (Wang et al., 2003). For simulation purpose, overall reactions can be simplified as suggested by Mitchell (1990): C + O2 CO2 + heat [65°- 94°C] (1) CO2 + C 2CO + heat [100°- 150°C] (2) Stoichiometric reactions (1) and (2) show that 2.66 grams of oxygen are required to oxidize 1.0 gram of carbon. Once the temperature exceeds 100°C, the reaction has little chance to stop. The crucial step in reducing spontaneous combustion risk is locating the ignition point of spontaneous combustion (hot spot). This information is useful in the effort of developing a preventive method effectively. In this study, this was conducted based on the best gathered information: ventilation surveys conducted in an existing longwall mine located in the western US; the laboratory experiments performed on a physical gob model; and the CFD models to investigate the flow behavior in the gob, the oxidation of coal, and heat transfer phenomena. The hot spot location is determined as a function of oxygen concentration and gob temperature. The critical values are the following: 5% (by volume) for oxygen and 100°C for gob temperature. CHARACTERISTICS OF GOB MATERIAL AND MODELS Permeability Concept Permeability is one of the key parameters in any study dealing with porous medium. This is determined by particle size and the ratio of void volume to the total volume (porosity) of porous medium. This relationship is known as the Carman-Kozeny equation (Scheidegger, 1957) and is given mathematically as [ ] where k* = theoretical specific permeability, m2 dm = the mean particle size, m n = porosity However, permeability tests with various particle sizes were carried out at the University of Utah?s ventilation model (Figure 1) to verify this relationship. These tests indicated a necessity to modify this theoretical relationship. A factor of 0.898 was obtained to make up the different between theoretical permeability and the experimental one. The modified version of (3) is given by: [ ] This equation is used to determine permeability of the simulated mine gob throughout the study. [ ] Mine Gob Material The gob is represented by zones filled with material of given size distribution. For physical measurements and CFD simulations, the material used were crushed rock and coal. Due to compaction, three different characteristics of gob material were used in the model: unconsolidated, semiconsolidated, and consolidated. In the real condition, the largest coal-rock particles are more likely to be located in the area behind the shields. This material is freshly broken and unconsolidated. The size of these broken particles is based on a study carried out by analyzing images taken from the area behind the shields in three coal mines in the United States (Pappas and Mark, 1993). The results have shown that the mean size of materials behind the shields is about 1.22 m. The permeability associated with this size was then obtained from (4) and permeability tests. This was 4.68 x10-7 m2. For semiconsolidated and consolidated zones, due to the lack of experimental data, these were determined through CFD simulations. These simulations were designed by assigning a permeability of 3.15 x 10-8 m2 for consolidated zone and 7.98 x 10-9 m2 to the other zone. The airflow pattern in the gob assigned with given permeabilities was expected to follow the experimental distribution (Brunner, 1985). The mean particle sizes for the semi-consolidated and consolidated zones were 0.02 and 0.006 m, respectively; smaller than those of the unconsolidated zone. Figure 2 shows the 3D view of permeability changes in the gob.

Jan 1, 2009

By Joe Roby

Rolling disc cutters are the business end of hard-rock and mixed-face tunnel boring machines (TBMs). Since their first successful employment on a TBM more than 50 years ago, disc cutter technology has constantly evolved, allowing modern TBMs to excavate very hard and abrasive rock efficiently. Disc cutters have also been employed successfully on earth pressure balance (EPB) and slurry machines, cutting rock under water and ground pressure. The range of materials excavated by machines today is broader, excavation rates are higher and cutter costs are lower than ever before, proving the value of investment in cutter development. Disc cutters are used on a range of tunneling equipment, from pipe jacked slurry micro-TBMs less than 1 m (3.3 ft) in diameter to 15 m (49 ft) in diameter, hard-rock boring TBMs. The geological conditions under which they are employed range from sands and gravels, with several bar of water pressure, to extremely hard, massive rock with UCS up to 420 MPa. Regardless of the type of machine or geology, one thing remains constant: changing worn out cutters is costly. When cutters must be changed in the middle of a tunnel, the contractor incurs the cost of downtime as well as the cost of refurbishing or replacing the cutters. In the case of catastrophic cutter failures, the project can be stopped completely for long periods, and the costs mount up rapidly while tunnel production is at a stand still. Catastrophic cutter failures include the occasion when hard-rock disc cutters are failed in groups (called a ?wipeout? phenomenon) and the operator fails to stop the machine, which results in severe damage to the cutterhead. The cause of this can be either an undetected failed cutter propagating damage to surrounding cutters or operator error in steering. Regardless of the cause, the resulting damage can take days or even weeks to repair. When cutters fail prematurely on EPB or slurry machines in certain geological conditions, it is impossible to evacuate the chamber and, therefore, impossible to get [ ] into the chamber to change the cutters. The solution is frequently an unplanned intervention shaft that must be sunk at great cost. The importance of being able to predict cutter life accurately, and in all geological conditions, cannot be overemphasized. For these reasons, cutters have been developed for specific machine types and sizes, as well as for specific geological operating conditions. Different size cutters are required for different size machines. For example, while 483- and 508-mm (19- and 20-in.) cutters are used on large diameter TBMs, it is impossible to employ cutters so large on a small micro-TBM. Also, the cutter must be designed for the specific geological conditions under which it will be operating. The disc cutter rings required to bore extremely hard rock are the most expensive of cutter rings. However, they provide little advantage in weaker rock formations where less expensive rings will do the job as well. It is important to choose the correct cutter for the machine and the geological conditions in order to get the best balance of cost and risk.

Jan 1, 2009

By A. C. Legg, M. T. Malema

Keywords: Pyrometallurgy, flash, smelting, furnace, nickel, copper, Botswana BCL Limited operates the only Flash Smelting Furnace on the African continent, at Selebi-Phikwe, situated in north-eastern Botswana. The furnace was commissioned in 1973 and produces a high-grade sulphide matte containing nickel, copper and cobalt, which is shipped to refineries in Zimbabwe and Norway for further processing. This paper outlines the recent process improvements and plant expansions that have resulted in a capacity increase from 40 000 to 60 000 tons per annum of metal production.

Jan 1, 2006

By M. Drozd, M. Somppi, Z. Yamak

"Granular activated carbon is widely used in the gold mining industry for gold extraction from aurocyanide-containing slurry. The extraction process can employ carbon as part of the gold recovery operation for various circuit designs including carbon-in-pulp (CIP), carbon-in-column (CIC), and carbon-in-leach (CIL) circuits. However, various qualities of activated carbon exist and the durability of the material can change significantly from source to source. Handling this expensive reagent can pose challenges when fine carbon losses occur in the CIP, CIC, and CIL circuits. Additionally, any gold that is adsorbed to these carbon fines can also be lost by various mechanisms including carbon attrition and fracturing, inefficient coarse-fine separation, and poor carbon fines capture. With the above in mind, a series of tests were conducted to investigate standardizing simple laboratory methods for testing carbon quality. A standard procedure beyond or enhancing ASTM methods can serve as additional starting points for evaluating carbon specific to the gold industry. The investigation focused on size distribution, flatness, wet attrition resistance, surface area, gold kinetics, and gold loading. The final results provided information to generate an equation for ranking carbon samples to plant performance.IntroductionActivated carbon for use in gold extraction is amongst the most commonly found processes in modern gold-milling plants. Although gold losses are inevitable in the milling process, operators often focus on the importance of maximizing gold recovery. One focus area that is important in gold recovery is optimizing the use of activated carbon and examining potential changes to realize improved economics for carbon-in-pulp (CIP), carbon-in-column (CIC), and carbon-in-leach (CIL) processes."

Jan 1, 2017

By S. A. H Tyrrell, R. Rotzinger

"Taseko Mines Limited recommenced operation of the Gibraltar copper-molybdenum mine in 2004 and embarked on series of development projects to modernize the mine and increase throughput capacity. The Gibraltar Development Plan 3 (GDP3) project included a second stand-alone copper-molybdenum concentrator, a stand-alone molybdenum separation plant and associated services. Ausenco was contracted to commence the detailed design in 2Q2011, commissioning commenced in 4Q2012 and first ore was processed in 1Q2013. This paper discusses the design, implementation and commissioning of the GDP3 Concentrator.INTRODUCTION The Gibraltar Mine (Gibraltar) property is located in the Cariboo Region of south central British Columbia, approximately 360 kilometers northeast of Vancouver. The site is accessed via an 18 kilometer paved road from the town of McLeese Lake, which is located on the province’s main north-south highway between the cities of Williams Lake and Quesnel. Gibraltar was originally developed by Placer Development and commenced production in 1972. Gibraltar was acquired, while on care and maintenance, by Taseko Mines Limited (Taseko) in 1999 during a period of unusually low metal prices. Taseko recommenced operations at Gibraltar in 2004, and embarked on series of development projects to modernize the mine and increase throughput capacity. The 1972 Gibraltar process flowsheet incorporated two-stage crushing followed by rod and ball milling and flotation. In 2008 Gibraltar completed their first development stage that included the installation of a SAG mill, converting the existing rod mills to ball mills and adding additional flotation capacity. Gibraltar continued development with the installation of an in-pit primary crusher, expansion of the bulk regrind circuit, expansion of the cleaner flotation circuit, replacement of the tailings pumping system and replacement of the concentrate filters and dryer."

Jan 1, 2014

By Gaulton R

The Commission maintains a data base first computerised in 1957 and containing information from over 18 000 boreholes located mainly in the Latrobe Valley area. Some of the borehole derived information dates from the turn of the century, although most of it has been collected as a result of Commission investigations conducted over the past 40 years. The Resource Data Base contains not only information gained directly from the drilling itself but includes the results of chemical and physical testing on many hundreds of thousands of samples retrieved from boreholes. Replacement value of the Resource Data Base has been estimated to be around $200 million, and it's integrity and accessibility is clearly of paramount importance, not only to existing mining and power generating functions but to the future evaluation and use of the vast brown coal Resource.

Jan 1, 1992

By Jack E. Bailey

From a start in 1957, milling 3,000 TPD of copper ore, the Cyprus Pima Mining Company, which is located 20 miles south of Tucson, has expanded its operations four times, and is currently treating over 55,000 TPD of ore.1 The most recent expansion added a 14,000 TPD semi-autogenous concentrator to the operation. Shortly after startup in January, 1972, the semi -autogenous plant exceeded the designed tonnage rate.2 Ore from the mine is delivered to a 54 inch gyratory crusher and is reduced to a minus 8 inch product. From the crusher the ore is stacked on a surge pile and then reclaimed and fed to the two parallel grinding circuits. A 28 ft. diameter by 12 ft. long semi-autogenous mill and a 16.5 ft. diameter by 19 ft. long ball mill make up each grinding circuit. The semi-autogenous mills are in closed circuit with screens and the ball mill circuits are closed with hydro- cyclones. Rougher flotation follows the grinding operation. Concentrate regrinding and cleaner flotation is performed in Mill I, the older mill. The generalized flowsheet is shown in Figure 1.

Jan 1, 1975

By Mathew Avery, Colby Barrett

In November 2016, a 7.8-magnitude earthquake caused more than 100,000 cubic meters of rock and debris to tumble down and bury the transportation infrastructure below Ohau Point in New Zealand. All routes between the northern and southern portions of South Island were severed. To undertake the rebuilding efforts, the New Zealand government established North Canterbury Transport Infrastructure Recovery (NCTIR) – an alliance to repair the road and rail networks between Picton and Christchurch. NCTIR’s aim was to have the road open by Christmas 2017, but work also continued into 2018 to complete a number of safety improvements along the highway. The presentation will review the initial remediation attempts, the necessity to re-evaluate the original efforts, the need to incorporate techniques that were “outside-the-box” and different from traditional methods, and the logistics and planning that were required to complete the work safely. INTRODUCTION Shortly after midnight on Monday 14th November, a magnitude 7.8 earthquake struck the small South Island settlement of Waiau in North Canterbury, 100km north of Christchurch. The quake was the largest in New Zealand since the magnitude 7.8 Dusky Sound earthquake in 2009 (GNS 2016). While shaking was widespread with over 15,000 recorded ‘felt reports,’ the worst shaking occurred about 50 seconds after the quake rupturing started. The energy of the tremor progressed north over several minutes with surface rupture recorded on a total of 21 faults. The length of all the fault ruptures combined was close to 100km (GNS 2016). The degree of ground shaking was high, recorded as Modified Mercalli Intensity Scale (MMI) Level 8 – Severe. This level of shaking causes considerable damage in ordinary buildings with partial collapse including fall of chimneys, factory stacks, columns, monuments, walls, etc. At a human level, people experience difficulty standing; while furniture and appliances shift. Geologically, the damage depends on the geological setting. In Kaikoura and the surrounding area, especially the coastal highway State Highway 1, north and south of the town, many of the slopes consist of over steepened weathered, fractured greywacke. During the quake, the shaking caused significant and widespread damage with a total of 26 major slips (and many smaller slips); closing both State Highway 1 and the Main North Rail Line between Picton and Christchurch.

Jan 1, 2018

By B C. Sarkar, D Ray, V Sarkar, P K. Sinha, R K. Singh, A Kumar

Estimation of iron ore grade distribution has been attempted using geostatistics and artificial neural network (ANN) based models for a banded iron formation (BIF) type iron ore deposit of East India. Geologically, it is a supergene iron ore formation through an enrichment process of parent material (BIFs) by gradual leaching of impurities, namely SiO2 and Al2O3. As the degree of leaching differs spatially, distribution of Fe varies depending on local controls of ore formation. While geostatistical methods work on models based on spatial correlations of mineralisation parameters, ANN models a non-linear distribution and is widely used as a global approximator. Spatial distribution modelling of Fe values has been carried out employing ordinary kriging (OK) and sequential Gaussian simulation (SGS). Non-linear modelling has been performed using hybrid neural networks (HNN) that provided an alternative option for grade estimation. Classical statistical modelling revealed a negatively skewed 3-parameter lognormal Fe population having a mean of 60.49 per cent with a skewness of -0.08 and a kurtosis of 2.60. Semi-variography and geostatistical estimation employing OK procedure led to block-by-block kriged estimates with associated kriging variance, which provided an overall mean kriged estimate and kriging variance of Fe as 60.29 per cent and 7.76 (per cent)2 respectively. SGS procedure revealed a mean of 60.71 per cent and a standard deviation of 5.89 per cent with minimum and maximum values as 45.12 per cent and 68.70 per cent respectively in respect of Fe. HNN model provided values of mean squared error (MSE) and correlation coefficient (R2) as 12.33 and 0.61 respectively, thereby exhibiting the robustness of the model to sustain maximum heterogeneity within the target domain. A comparison of Fe grade estimates obtained by kriging, simulation and HNN models more or less maintain similar distribution patterns that are close to the sample data. It is thus opined to carry out the grade prediction exercise employing all these three modelling approaches to obtain alternative scenarios for improved mine planning and design exercises. The techniques provide a foundation for cogent grade modelling leading to improved iron ore reserve estimation procedures.CITATION:Sarkar, B C, Singh, R K, Ray, D, Kumar, A, Sinha, P K and Sarkar, V, 2017. Iron ore grade modelling using geostatistics and artificial neural networks, in Proceedings Iron Ore 2017, pp 431–438 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

By Trevor Rangasamy

Technically based guidelines for managing rock related risks associated with strip mines in South African collieries are currently vague and are to a large extent based on experience and trial and error. To date, no mine specific or universal design charts exist within the South African industry. Quantification is done on a reactive basis i.e. following reported instabilities. The conditions under which, proactive tools to manage instabilities would be practically applicable are: The design of simple easy to use charts. Designing the charts for a target audience that is primarily production management. Well organized, logical set-up of the charts that will allow for easy future refinements. Allowing for variations in rock and soil properties within the charts that would account for colliery specific and region specific variability. The factors selected for in depth investigation that were deemed to have either individually or in combination, a primary influence on the stability of strip mines are (in no order of priority): Appropriate batter angles for ?soft? materials. Surcharge loading (spoils, dumps, topographical highs) on the highwall side. Influence of the thickness of stratification on stability. Influence of moisture content on stability. Influence of the changes to the coal seam dip angle on stability. Influence of persistent ?strip long axis? parallel faults/joints on stability. Spoil stability for variations in material type. This paper provides details of the work undertaken, using principally numerical modelling techniques, to produce design charts that will allow for proactive management of rock-related instabilities on strip mines.

Jan 1, 2008

By John Russell

"Faster, safer, grinding mill relining offers the obvious attractions of reduced mill relining times with the knock-on incentive to reduce all other maintenance down-time, opening the possibility for increased annual plant utilisation.Leading edge liner handler technologies by RME, focused toward even shorter relines for very large SAG and AG mills, have recently split into two quite distinct paths:•,?Single SAG/AG Mill Relining Machines (MRM’s) of very large capacity, such as Newcrest Cadia’s 40’ SAG MRM of 6000 kg capacity and P.T. Freeport’s 38’/34’ SAG MRM of 7250 kg capacity;•,?Twin-machine SAG/AG Mill Relining Systems of moderate to large capacity such as BHP Billiton’s Escondida Laguna Seca 38’ SAG MRM’s, each featuring 4500 kg capacity.RME, its customers noted above and their liner suppliers are providing leadership in this important annual utilisation advance.Once the 24 hour mill shell reline becomes reality, liners designed for maximum profit, not maximum life, albeit exchanged more frequently, become possible too.INTRODUCTIONAt most mineral grinding plants, mill relining activities dictate each maintenance shutdown’s critical path. When annualized, mill relining remains the largest, single influence in defining plant availability.Because of restricted access, both inside and outside the mill, increased application of manual labour has no effect on reducing relining downtime. The only answer for RME has been to develop and apply new relining technologies now known as RME’s Mill Relining System.RME’s primary focus for over twenty years has been the reduction of grinding mill relining times while simultaneously providing a marked improvement in relining safety. Our original (1990) target was to halve traditional (pre-RME) reline times and to eliminate accidents and injury by mechanizing all physical work."

Jan 1, 2008

By N. Sano

The chromium oxide in stainless steel slags usually exists as Cr2O3. However, there is a possibility that it is oxidized to hazardous CrO3 after disposal or during use. Therefore, some technical developments to reduce chromium oxide by using a smelting reduction technique have been reported recently. This paper describes the fundamentals of the technique which have been applied to ferrochromium production and reviews the recent results in literature.

Jan 1, 2004

By Chen Xia, G. Deschenes, Wan-Tai Yen

"Ammonium thiosulphate is a potential replacement for the cyanide in gold leaching. However, the high reagent consumption hinders its practical applications. In order to reduce the thiosulphate consumption, chemical additives were used. The gold leaching tests were conducted on a sulphide copper-bearing gold ore with 0.3M thiosulphate, 0.03M copper sulphate and 3M ammonium hydroxide solution at 44% solids and pH 10.2. Sulphate, sulphite, amino acid and chelating agents were initially used as additives. It was found that chelating agents (EDTA and NTA) could reduce the thiosulphate consumption from 30 kg/t to 17.8 kg/t with some increase of gold extraction. The improvement from other additives was not significant. Other approaches were also attempted to reduce thiosulphate consumption with some success. The thiosulphate consumption could be reduced to about 17 kg/t by reducing copper sulphate concentration fiom 0.03M to 0.0015M; adding 0.04M of EDTA; replacing copper sulphate with 0.001M of nickel sulphate. The increase of pulp density fkom 44% to 60% solids could reduce the thiosulphate consumption to 17 kg/t as well. Limiting the dissolved oxygen content by sparging nitrogen reduced the thiosulphate consumption further to 10 kg/t. It could be concluded that the regulations of copper and oxygen supply in the lixiviant were the keys to reduce the reagent consumption.INTRODUCTIONCyanidation, as the predominant gold leaching method, has been used in mining industry for more than a century. The low reagent requirement and the high performance are the reasons for its great success. However, cyanidation cannot deal with ores containing large amount of copper, nickel, arsenic and zinc minerals. This type of ore may consume more reagents and result in a poor gold extraction. On the other hand, the carbonaceous materials may preg-rob the gold from pregnant solution and make cyanidation a problem leaching method. Further more, the high toxicity of cyanide represents another challenge to modern gold mines. In the beginning of the year 2000, an environmental disaster caused by an accidental spill in Yugoslavia was broadcasted world-widely (Gurevitch 2000). Gold mine operators are facing higher social pressures and increasing costs on environmental protections. The usage of cyanide in mining industry has been banned in some countries and locations such as Czech Republic, Japan, Turkey, California, Colorado, Wisconsin and Montana States. Thus, a non-cyanide lixiviant is required."

Jan 1, 2002

By Graham A. Davis, Michael Samis

The dramatic fluctuation in metal and energy prices in the last few years is causing mining companies to consider alternative analytical methods for mine design and valuation. The goal is to improve insight into unforeseen changes in the economic and project environment and to design projects that take advantage of management's ability to respond to these changes. Such flexibly managed projects will contain value that cannot be measured using traditional engineering economic approaches, further motivating the used of alternative, modern valuation methods. This paper considers the valuation of an uneconomic low-grade resource at an open pit gold mine using two of the most popular modern valuation methods, Dynamic Discounted Cash Flow and Real Options. Our analysis demonstrates that the conventional valuation method may incorrectly represent the value and risk attached to mining this type of resource because it is unable to recognize management's ability to respond to future events in a manner that minimizes losses and exploits advantageous business conditions. The results of this paper highlight the importance of recognizing flexibility and business uncertainty in both mine design and valuation.

Jan 1, 2003

By P. Fossi

This report describes a pilot-scale study of a hydrometallurgical process to treat high-grade nickel matte or a nickel-cobalt sulphide precipitate to produce high-purity electrolytic nickel. The first step of the process is the oxidative leaching of the nickel-bearing material by nitric acid; the nitrous gases generated during leaching are reconverted to nitric acid. The crude nickel sulphate nitrate solution is purified and is then subjected to a denitrification step. Nitrate separation is effected either by the crystallization of nickel sulphate or by precipitation of basic nickel carbonate. The electrolyte is prepared from the nitrate-free material and nickel is then recovered by electrolysis of the sulphate solution using insoluble anodes. Some technical and economic problems encountered with this processing route are discussed and the method is compared to the "all-chloride" process developed by Société Le Nickel.

Jan 1, 1977

By C Hall, J L. Mauk

New 40Ar/39Ar dates of vein and wall rock adularia from seven epithermal deposits and one porphyry copper prospect help constrain the ages of mineralisation in the Hauraki Goldfield. Vein adularia from the Paritu region yielded a single 40Ar/39Ar plateau date of 16.315 ¦ 0.066 Ma. Five 40Ar/39Ar plateau dates of three samples of vein adularia from the Opitonui deposit overlap within error to yield a preferred age of 13.149 ¦ 0.016 Ma. Two samples of vein adularia from the Night Reef at the Broken Hills deposit yield exceptionally flat 40Ar/39Ar plateaus, and all dates overlap within error to provide an age of 7.121 ¦ 0.010 Ma. Two 40Ar/39Ar plateau dates of one sample of vein adularia from the Wharekirauponga deposit overlap within error to provide an age of 6.318 ¦ 0.061 Ma. Four 40Ar/39Ar plateau dates of three samples of vein adularia from the Maratoto deposit overlap within error to provide an age of 6.411 ¦ 0.022 Ma. Two samples of adularia from altered wall rock at the Sovereign deposit yield three 40Ar/39Ar plateau dates that range from 6.547 ¦ 0.021 to 6.806 ¦ 0.028 Ma. This suggests that mineralisation formed around 6.6 Ma, but the results are complicated because the adularia contains minor illite that post-dates adularia formation. Results from the Karangahake deposit are complex. Two samples of vein adularia yield three 40Ar/39Ar plateau dates that range from 6.116 ¦ 0.116 to 6.901 ¦ 0.101 Ma, whereas one sample of wall rock adularia that contains minor late illite yields two 40Ar/39Ar plateau dates that overlap within error to provide a date of 5.734 ¦ 0.041. The range of dates indicates that more work is necessary to understand the history of mineralisation at this deposit. Two samples of vein adularia from the Tui deposit yield four 40Ar/39Ar plateau dates that overlap within error to provide an age of 5.714 ¦ 0.017 Ma. These results confirm previous work that indicates that the ages of epithermal deposits young southward in the Hauraki Goldfield. These results are also consistent with previously postulated different ages for deposits hosted in Kuaotunu Subgroup rocks in the northern goldfield (>10 Ma), and deposits hosted in Waiwawa Subgroup rocks in the southern goldfield (7.1 to 5.7 Ma).

Jan 1, 2005

This paper sets out to examine the possible relationships between three factors, they are, the rate of growth of world production of minerals, the growth of Australian production and the growth of scientific and engineering knowledge.The first two are quantified, in terms of tonnes and value. The third, an abstract concept "knowledge", requires to be quantified, in terms of manpower.A mathematical model, which has general application, was derived to demonstrate certain axiomatic premises and to define, more precisely, basic concepts related to the philosophy of knowledge.Finally a summary is made of education-cum-industry issues.

Jan 1, 1981

Valuers of mineral properties should be aware that accountants are a major user of their opinions of value. Reliance by accountants can range from obtaining comfort that book values in balance sheets are not overstated to establishing values by which to assess the fairness of a takeover offer. Accountants using mineral property valuations clearly can't be expected to have the same level of expertise as the valuer and ""second guess"" the valuation. But they do have an obligation, under the ASC guidelines and their own professional standards, to'check the logic and reasonableness of the valuation before being able to rely on it. This means reviewing the valuation report for general sensibility, discussing the methodologies and assumptions with the valuer and being satisfied as to his competence and independence. This paper outlines how accountants use mineral property valuations, the level of scrutiny applied to such valuations and the problems commonly encountered in practice by accountants seeking to rely on them. The paper also explains how values of-mineral properties are established in company balance sheets and why these often differ, quite validly, from fair market values.

Jan 1, 1994