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By S Robotham, P C. Stewart, K Maher

A data science based dilution study was conducted to identify the causes of dilution at South32’s Cannington Mine (Cannington), a large open stoping operation. In terms of industry benchmarks for unplanned dilution, Cannington is already performing very well, and so opportunities for improvement are more difficult to identify. The study is based upon back analysis of nearly 150 case studies from the mines reconciliation database, and has quantified the relative effect and significance levels for the following dilution parameters; time until paste filling, faulting and cable bolt reinforcement. Specifically, significant contributions were associated with faults and stope open duration. Quantifying the effect of faults on stability graph accuracy enabled a site specific adjustment to improve the predictive ability of the Cannington stability graph (unstable stope specificity improved from 63 per cent to 74 per cent). Despite Cannington’s already high levels of stope performance compared to industry benchmarks, data science methods have identified a complex interaction between identified dilution factors (ie stope open duration, presence of faults and cablebolting). Specifically, a third order Factorial ANOVA (Analysis of Variance) provided evidence that stope open duration reduces the effectiveness of cable bolts in faulted ground (average dilution = 2.1 m, p-value = 0.13). These results are consistent with site observations of a relationship between the stope open duration and the extent of failure when unravelling on a fault, even when cables are present.CITATION:Maher, K, Stewart, P C and Robotham, S, 2016. A data science approach to identifying and quantifying causes of dilution at Cannington Mine, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 491–500 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

Concentrations of bright green material in the "Golden Mile" of Kalgoorlie, which carry particularly high gold values, are known locally as "green leader".This assemblage consists largely of quartz, sericitic muscovite, carbonates, and pyrite. Some of the sericitic muscovite contains vanadium, giving the rock its green colour. Vanadium oxides, including nolanite, are also present. Tellurides identified in this study are calaverite, altaite, petzite, melonite, and coloradoite. Native gold was found in most of the samples; some of the gold, at least, has been formed from the decomposition of gold tellurides.It is suggested that mineralizing solutions carrying gold and tellurium leached vanadium from the host rocks and subsequently deposited gold tellurides and vanadium-bearing minerals in the "green leader".

Jan 1, 1977

By B. Simonian

"The Mount Milligan mine, located in central British Columbia, began operation in September 2013 with a design nameplate capacity of 62,500 dry metric tonnes per day, using a conventional SABC grinding circuit followed by bulk sulphide flotation. Over the last few years, many changes have been implemented in order to debottleneck and optimize the comminution circuit to meet and exceed the nameplate design capacity as well as improve the metal recovery of the flotation circuit. This paper discusses changes, grouped into mechanical modifications, projects, and operational changes. The implementation of these changes as well as ongoing challenges and potential solutions will also be discussed.INTRODUCTION The Mount Milligan mine began production in the fall of 2013. It is a conventional truck-shovel open pit gold and copper mine and concentrator with a 62,500 tpd design capacity copper-gold flotation processing plant. The average head grades for the life of mine are 0.18% Cu and 0.39 g/t Au. At full capacity, the mine produces an average 400 dry tonnes per day of copper-gold concentrate containing approximately 21.5% Cu and 50-60 g/t Au. Current mine life is 21 years. This paper will discuss a selection of the operational, engineering, and administrative changes that have been made at the site since the publication of An Update on Reaching Nameplate Throughput and Continuous Improvement at Mount Milligan (Yue, 2017). PLANT FLOWSHEET CONFIGURATION Run-of-mine ore is crushed through a 60-110 gyratory crusher to a top size of seven inches. Up to 100% of the gyratory crusher product is diverted to the secondary crushing circuit by a moveable gate. The material diverted to the secondary crushing circuit is fed to a vibrating double-deck screen. The screen oversize is placed on a surge pile and reclaimed by four vibrating feeders to feed two 1250 kW standard head cone crushers in open circuit. The screen undersize is combined with the crusher discharge and returns to the conveyor feeding the coarse ore stockpile along with primary crusher product that by-passes the diverter gate."

Jan 1, 2019

By Patricio Vergara Lara, Jorman Barrera

The selection of the production level elevation is one of the most critical strategic decisions faced by large-scale block or panel caving projects. This elevation largely affects project timelines, capital expenditures, resource recovery, and may even constrain the mine’s maximum production capacity. Over the years, many computational tools have been developed to assist in this decision, most of them based on the same principle of maximizing the project value given its production level elevation. However, for large and inclined deposits, or those extending significantly in the vertical dimension and where extraction cannot be achieved from a single level, traditional selection techniques involving multiple production horizons often lead to suboptimal results. This paper proposes a methodology ensuring that the selection of production level elevations in a multiple horizons (“lifts”) case is globally optimized. Input parameters include economic and cost variables, maximum extractable column height, and deepening costs. The methodology is demonstrated through a real case study to illustrate its potential impact on project value.

Feb 22, 2026

By D. W. Holliday

The major deposits of anhydrite and secondary gypsum in Jamaica occur discontinuously at a single horizon within the Lower Eocene succession of southeast Jamaica. They are not intruded from other horizons of different age, as has previously been supposed. The gypsum and anhydrite occur in a regressive sequence at the junction between marine (Richmond Formation, Halberstadt Volcanics, Halberstadt Limestone) and overlying non-marine (Wagwater Formation) sediments. They are thought to have been formed in an arid supratidal (sabkha) environment. A variety of 'nodular' structures occurs in the gypsum and anhydrite which can be related to early diagenetic processes and compaction effects. All gradations exist between single isolated anhydrite nodules and structureless anhydrite due to closer packing and coalescence of nodules. The structures and textures in the anhydrite are thought to be consistent with penecontemporaneous formation from gypsum crystals rarely more than 2 cm in diameter. Single anhydrite nodules are not usually larger than 2 cm; nodules of greater size are compound, due to the coalescence of several smaller nodules. Some large compound nodules may have developed from single sand crystals or rosettes. Compaction of anhydrite appears to have been largely mechanical, but some rocks have an aligned fabric, suggesting later flowage

Dec 1, 1971

By M. G. Lipsett

"The Mining Information Technology User Group considers information to be a strategic enabler to improved business performance in mining. This ad hoc group has been meeting since 2000 to share knowledge and leverage their efforts in effectively using IT in mining. In this paper, core issues to mining IT are reviewed, along with initiatives to engage industry professionals and decision-makers in the potential for improved use of information in their operations. Relevance of IT to Canadian MiningMining operations always drive to improve their performance, and so far they have succeeded by using different technologies. While economies of scale have come in fragmentation and material handling, and energy usage has improved through conservation and design of integrated mineral processing plants, these systems have the net result of tightening operational constraints, through reduced excess capacity. There are higher opportunity costs of lost production when integrated, large-scale systems are operating in a sub-optimal way.The problem is that it is profoundly difficult to run a mine in an optimal way. Optimization requires a defined objective function (e.g., profit), which is hard to pose, let alone solve, because so many inter-related aspects of an operation affect the bottom line. Even for a small, well-defined process state, it is difficult to measure all the pertinent information with sufficient accuracy to ensure that the system is performing well in real time.The art of mining remains making good decisions with uncertainty: in the orebody delineation; in the allocation of resources; in the efficiency of the operation; in the reliability of components and systems; and in responding to the downstream market. As well, individual decision-makers need to know how their decisions affect other processes and systems, so that improvements in one part of an operation do not result in such degraded performance in the rest of the plant that overall profitability goes down."

Jan 1, 2003

Hard rock TBMs are very often open machine layouts, with just a part roof shield over the front of the machine. The whole system is designed for fast face advance and the economy of the project often depends entirely on the rate of advance. Typically, hard rock conditions will demonstrate stable and good ground for a major part of the tunnel. However, short sections of crushed shear zones with clay and gouge material may cause serious time delays. Spiling rock bolting has been proven to be very efficient under such circumstances, provided the fully grouted rebar bolts can be placed in a controlled and practical manner.Environmental restrictions are becoming more and more a part of tunnel excavation projects. Even hard rock TBM tunnelling may require a strict ground water control, because of consequences on the ground surface from lowering the ground water level. Such consequences may be surface settlement, negative influence on ground water resources or damage to vegetation.Pressure injection ahead of the advancing tunnel face (in this paper named pre-injection), using microfine cement and chemical grouts, is today a proven technology. As examples can be mentioned (Garshol, 1997) about 100 km of pre-injected tunnels below Oslo City and about the same length of subsea tunnelling elsewhere in Norway. Examples from other parts of the world are also available, like below Stockholm, Sweden, the Inland Feeder project and Hollywood Metro both in Southern California.To take advantage of spiling rockbolts and pre-injection, it is an obvious prerequisite to be able to drill the necessary boreholes in the right positions and at the correct angle. In drill and blast excavation this is simple, but in TBM projects it has repeatedly turned out to be difficult. More than once owners have accepted bids containing reassurance from the contractor that the drilling method will be sorted out later. Experience shows that if ælaterÆ means after start of the TBM operation, it is often too late.

Jan 1, 1999

By John G. Payne

The Pebble deposit contains three contiguous zones. In the Pebble West zone, granodiorite stocks intrude gently warped volcanosedimentary rocks that had been intruded by diorite sills and two felsic intrusion breccias. In the Central zone, volcanosedimentary rocks are cut by diorite and granodiorite sills; the latter are genetically associated with the granodiorite stocks. In these zones, widespread potassic alteration consists of brown biotite and K-feldspar, with abundant quartz-carbonate-sulphide vein stockworks. Chalcopyrite/pyrite ratios range from 2/1 to 1/20. Diorite hosts the highest-grade mineralization, with much of it being disseminated. Patchy, later phyllic alteration is moderate and commonly depleted in copper. The Pebble East zone occupies a granodiorite stock and intruded host rocks below an eastwardly thickening wedge of barren Tertiary volcanic and sedimentary rocks up to 600 m thick. Brown biotite alteration and copper mineralization are strong, with chalcopyrite/pyrite ratios commonly over 1/1. The zone extends to a depth of at least 1200 m below surface; some drill holes bottomed in strong mineralization. Later, weak to moderate sericitic alteration removed some chalcopyrite and locally remobilized gold. The Pebble deposit has a Measured/Indicated resource of 3 billion tonnes grading 0.28% Cu, 0.32 g/t Au and 0.015% Mo (= 0.5% copper-equivalent) and an Inferred resource of 1.13 billion tonnes grading 0.30 g/t Au, 0.27% Cu and 0.015% Mo (0.55% copper-equivalent). The Pebble East zone has an Inferred resource of 2.13 billion tonnes grading 0.55% Cu, 0.34 g/t Au, and 0.037% Mo (0.97% copper-equivalent). The Pebble East zone is open to the east and to a lesser extent to the north, south, and to depth.

May 1, 2006

By Amalia Giannakou

This paper utilises 3-D finite-element modeling to study the static and dynamic lateral harmonic response of inclined piles, in a homogenous and a non-homogenous soil, modeled as a linear hysteretic medium. The results of a comprehensive parameter study are presented in the form of dimensionless plots of: (a) static and dynamic swaying and rocking stiffnesses and damping ratios, and (b) distribution of internal forces (bending moment, shear and axial force) with depth. These plots can be readily utilised in practical applications and they offer insight on the role of the inclination in the response of flexible piles.

Jan 1, 2006

By F. P. Haver

The Federal Bureau of Mines developed a 10,000-ampere cell to determine the economic feasibility of producing high-purity electrolytic titanium from mill scrap or offgrade sponge. Operating variables were investigated, matetial handling procedures were established, and the cost of electrorefining titanium was determined. The prototype cell produced premium-grade metal (<100 Brinell hardness number (B.h.n.)) from very impure Kroll sponge at a processing cost of less than $0.60 per pound. Refined metal was deposited at the cathode in the form of coarse (85 percent plus 34-mesh), hexagonal crystals of a titanium when the cell was operated an 850° C. A cathode current density of 1,000 amperes per square foot and a molten sodium chloride electrolyte containing approximately 5 percent by weight of soluble titanium were used. In continuous operation, at the optimum number of ampere-hours, (40,000), the current efficiency averaged slightly less than 60 percent. It was possible to supply large enough amounts of electric current for electrolysis to maintain the temperature of the bath, thus reducing overall power consumption to less than 5 kilowatt-hours per pound of metal refined.

Jan 1, 1961

By D. M. Wyslouzil

The fundamentals of dissolving silver and silver minerals are discussed in the following sections: 1) Basic chemical reactions in the silver system 2) Leaching of pure silver minerals 3) Leaching of silver-bearing ores 4) Description of processes The basic chemical reactions of silver are an important guide to the mineral processor or the extractive metallurgist, particularly for systems in which complex formations are balanced against solubilities. The competition of other cations for free ligands can shift the equilibrium to the detriment of silver. Solubilities of pure silver minerals in specific solvents may not be a good indication of the solubility of silver minerals in ores. The abundance of silver minerals which usually appear in a typical ore precludes the selection of a single method that would result in the best extraction of silver from all minerals in the ore. Frequently, the recovery of silver from host minerals requires a destructive pretreatment such as intensive leaching, oxidation with or without elevated temperature and oxygen pressure, or roasting. The transformation of the host or the silver minerals does not always result in a leachable species? and this requires additional treatment before the precious metal can be recovered.

Jan 1, 1990

By Carvalho A. Felipe de

The search for environmentally friendly and cost-effective methods for gold extraction has been undertaken in order to provide alternatives to the use of cyanide leaching in the gold industry. Among such alternative reagents is the combination of chloride and hypochlorite ions due to their oxidation of metallic gold and its complexation. Therefore, the current work aims at investigating gold leaching from an oxidized material by means of a solution comprising calcium hypochlorite and sodium chlorite. The sample used in the current research contained 66.1% Fe2O3, 20.4%SiO2, and 35.7 g/t of Au. In addition, XRD and SEM-EDS analysis revealed iron oxide, quartz and muscovite as the main crystalline phases in the sample. For comparison purposes, cyanidation of the oxidized material resulted in 90% gold extraction in 24 h. The reference conditions for the leaching tests with the chlorite-hypochlorite system were: particle size −37 μm, solid/liquid ratio of 10% (w/v), 10 g/L NaCl, 10 g/L Ca (OCl)2, pH 5–6, 3 h of leaching. The effects of the: (i) Ca(OCl)2 concentration, (ii) NaCl concentration, (iii) pHrange and (iv) the effect of pulp density in the extraction of gold were assessed. In addition, the stability of the hypochlorite solutions (OCl−) was investigated. The ideal concentrations of Ca(OCl)2 and NaCl were determined as 25 g/L and 15 g/L, respectively, whereas the optimal pH range was 4–6. Under these conditions, the final gold extraction reached approximately 84%. By varying the solids content of the pulp (10%, 20% and 30%), it was observed that the lower solids value resulted in greater gold extraction (84%). The degradation of the hypochlorite solution and the possible catalytic effect of the presence of the solid were evidenced. To sumup, in the experiment, high gold extractions were achieved in shorter periods of time (3 h), as compared to the cyanidation of the same sample. On the other hand, itwas necessary to use significant amounts of the reagent due to hypochlorite instability at the pH conditions recommended for high gold extractions.

By K. Lim, J. Sobil, L. Galarza Castillo, F. Tozlu

"For mining operations globally, the scale and criticality of the grinding mill drive has meant that selection of the correct mill drive technology is paramount. With so much depending on the successful implementation of the electrical drive system— safety operation, timely installation, availability, energy efficiency, minimum maintenance and extend its use as tool to assist during mill liner exchange—customers are justified in their scrutiny and slow adoption of new drive technology. From the first installation of low, fixed speed synchronous motors to the introduction of the current source and further voltage source variable frequency drive, innovations require long periods of time and mining customers willing to push the industry forward. In recent years, the variable frequency drive has been the stepping stone for a new generation of mill drive technology. In use for decades in other heavy industrial applications, the drives allow for system optimization in a myriad of ways. The most dramatic of changes is in the design of the motor. Where in the past motor design was bound by line frequency, today the designer can optimize machine size and, more importantly, eliminate the need entirely for the synchronous motor. Induction motor design has now progressed to the point where low speed operation is possible at efficiency levels nearing those of the synchronous design. Moreover, the simplicity of the squirrel cage design means that availability, reliability and spare part requirements are greatly improved. As the next step in the evolution of large pinion driven grinding mills, the low speed induction drive offers a new reliable and robust system to the market.IntroductionAs deposit grades continue to decline, new ways of processing ever greater concentrator throughputs have been required to maintain throughput efficient and profitable operations. Furthermore, the high risk nature of mining developments requires a constant search for safer, higher availability and more efficient equipment design. Limitation of downtime is a key factor in ensuring mining operations can be sustained through volatile market conditions. Often processing the entire mine throughput through just 1 or 2 lines, the crushing and grinding circuits are of particular criticality to the uptime and, thus, profitability of any mining operation."

Jan 1, 2016

By J. Stacey

This study was undertaken to generate a locally relevant guideline for the socio-economic aspects of mine closure for South African closure practitioners and decision makers. It should also represent a resource for interested and involved stakeholders. Mine closure, and its related post-closure social, environmental and economic characteristics, is part of the lifecycle of mining operations, given the finite nature of the ore resource. Many coal mines are anticipated to close in the next ten to fifteen years in Mpumalanga, as the epicentre of the industry shifts towards the Waterberg. Social aspects of mine closure have been historically underplayed, and under researched. Coaltech thus commissioned two studies, the first of which examined the closure dynamics of 36 mines in Mpumalanga and KwaZulu-Natal in 2006. A follow-up study (July 2007) focusing on Mpumalanga3,4 confirmed some of the results of the first study. It also contained broad guidelines for mine closure covering social and labour planning, job creation, the use of mine infrastructure, developmental incentives, and environmental rehabilitation.

Jan 1, 2010

By F. Ernest Brackett

Much new data on the flow of gases have been discovered by recent experiments by the United States Bureau of Mines and others. Although additional investigation is still desirable, the information now collected seems sufficient to attempt a tentative coördination of the friction factors. This study is confined entirely to fluid friction when the density throughout the duct is so nearly constant that it may be so considered. In the subsequent discussion the following notation is used: v = mean velocity of fluid in duct, feet per second. v&apos; = mean velocity of fluid in duct, meters per second. R = hydraulic radius of duct, feet. R&apos; = hydraulic radius of duct, meters. S = the hydraulic slope. P = pressure absorbed by friction, pounds per square foot. p = pressure absorbed by friction, pounds per square inch. p&apos; = pressure absorbed by friction, kilograms per square centimeter. A = sectional area of duct, square feet. O = perimeter of duct, feet. L = total length of duct, feet. L&apos; = total length of duct, meters. d = diameter of a cylindrical duct, inches. w = density of the fluid, pounda per cubic foot. g = specific gravity of the fluid compared with water = w % 62.4. C, k and B, coefficients for formulas using English measures. C&apos; = metric coefficient, v&apos; = C&apos;R&apos;S. n = coefficient of roughness for English or metric units. The basic formula for the flow of all fluids at constant density appears in the three following forms, which may be denominated: (a) The slope formula, v = CRS (b) The ventilation formula, " n = (c) The pipe formula, C is Kutter&apos;s coefficient, which is extensively applied to the flow of water in ducts and channels of all kinds; E is Atkinson&apos;s coefficient, used principally with air in ventilation problems; B is a coefficient conveniently adapted to studying the flow of water, gases and vapors in pipes. 312

Jan 1, 1927

By Gordon Revey

The challenge of managing blasting risk, types of risk, public perception, and a system for managing risk were covered in the first part of this article in the May/June 2000 issue of the Journal of Explosives Engineering. This final part of this two-part series provides a method for identifying specific blasting risks and reviews the importance of community relations.

Jan 1, 2001

By C. Siac

"It is a widely held view that in the 21st century, the mining industry will be dominated by a relatively small number of large companies, truly global in scope and internationally staffed (Gooding, 1997). Interestingly enough, most of the exploration and exploitation activities are being carried out by junior and major mining companies from common law jurisdictions in civil law countries, in particular in Latin America, where in recent years, the region has ranked first as a destination for exploration dollars, thanks to the reforms undertaken by some countries to amend their legal framework to attract foreign investment. From a lawyer’s point of view, this is where the main challenge lies: how to bridge the gap between the two legal systems.In order to avoid pitfalls, common law lawyers should be more diligent and proactive when dealing with civil law jurisdictions, given the different systems of law, standards, traditions and customs. Common law lawyers cannot be expected to be well-versed in civil law but they should not be completely unfamiliar with it either. A balance should be struck to ensure clients’ needs are professionally taken care of and that common law requirements and standards are met."

Jan 1, 2003

By L. E. Mordberg, E. N. Nesterova

Geochemical studies of the bauxite deposits show that the primary composition is gibbsitic. Some trace elements were relatively mobile during bauxite formation with the content of Th, U and Zr increasing and that of Rb, Ni and Co decreasing. Y and Pb were relatively immobile during lateritisation; their contents and ratios were inherited from different types of underlying rocks. The supply of Ni and Co took place at the diagenetic or epigenetic stages; their ultimate source was the crystalline basement in the region. The studies support the view that the bauxite was likely to have originated as a result of in-situ lateritisation and not through sedimentary transport and redeposition

Jun 18, 1905

By Andrew K. hg` Officer, A. L. Griffiths, J. Andrew Hindmarch, Gert Wittneben

Durban Harbour in South Africa is one of the continent’s busiest ports however the harbor entrance is currently too small for the new generation of container ships. There-fore it is to be deepened and widened by the National Ports Authority in the near future. An existing immersed tube tunnel carrying major sewer, water and other utilities under the harbor first needs to be removed and replaced by a deeper 4.4 m ID ser-vices tunnel. Construction of the deeper tunnel saw the first use of a slurry machine in South Africa and was required to negotiate grades up to 20%.

Jan 1, 2007

By Joseph A. DeCarlo, Eugene T. Sheri dan

The process of converting coal into coke is known as carbonization. Essentially, carbonization is destructive distillation in which the volatile matter of coal is released when coal is heated above its decomposition temperature and passes through a plastic stage before solidifying and forming coke. The mechanics of coke formation are complex, and the specific properties of certain coals that allow them to "cake" or form coke are still unknown. Coke is a hard, cellular, carbonaceous mass, formed by heating coal in the absence of air. It is porous, irregular in size and shape, and is similar in appearance to wood charcoal. Coke, however, is denser, more porous, and more abrasive than charcoal. Coke has been in use for several thousand years. History has recorded its use in Greece before the birth of Christ, and it is thought to have been used in China even before that time. It was not until the l6th century, however, that a method for coking was developed in Europe that subsequent ly evolved into the coking process that is known today. The coke industry in the United States began in the latter half of the 19th century when beehive-coke ovens were built to supply fuel for iron blast furnaces. Before this time, an iron industry fueled principally by charcoal had been in existence for several hundred years. With the depletion of timber resources, coke gradually replaced charcoal as blast-furnace fuel, and by 1875 more iron was made with coke than with any other fuel. Before 1893 a 11 coking was done in beehive ovens. These ovens are refractory kilns with arched roofs, sometimes joined together to form batteries.

Jul 1, 1964