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By E Skinner, G Lane

Project delivery strategies are determined by project context, business case requirements, scale of project, market conditions, ownerÆs capability and preference, contractorÆs experience, resource and skill availability, project location and type of project or facility.Successful project management requires the setting of coherent, effective and realistic business objectives and success criteria. The management team needs to lead, prioritise, resource, align, plan, track and communicate the processes that enable the realisation of these business objectives and success criteria.For any given project there is a unique balance between selection of a simple project delivery strategy that minimises interfaces and a complex strategy that necessarily engages the use of expertise and experience from numerous parties. The right balance for a project between these "book-end cases" is a function of how the project owner or managing contractor defines its objectives and chooses to manage risk, cost, schedule and project quality, in the context of the conditions affecting the project.Selection of an effective project delivery strategy requires a detailed appreciation of the range of external and internal environmental / shaping factors which influence the project. These factors include the project constraints imposed by the business, market or operational environments and the selection of an effective mix of primary (head contract), secondary (mix of vertical and horizontal packages) and tertiary (type of individual contract) strategies. The selection of the primary, secondary and tertiary strategies also shapes the project, the size and responsibilities of the ownerÆs team and the framework for delivery of outcomes.How does an owner or prime contractor decide? What is the optimum for a given project? Those with operations management backgrounds will have different views to those with project and construction management backgrounds.Is the simplest practical approach the best approach? Each and every interface needs to be managed.The owner will want a very experienced team that is engaged directly with the owner. The team will have a track record of successfully performing projects using contractors skilled in their respective disciplines and selected using competitive tendering on brilliantly designed scopes of work and good contract terms, with a fully defined project scope and appropriate schedule, for which all significant risks have been addressed and no scope changes are outstanding. Budget and schedule will be expected to have been developed with a full understanding of the scope, situation, risk management requirements etc.The problem is that these conditions rarely exist. The challenge in selection of a project delivery strategy is to resolve how best to address the inadequacies between the ideal case and the present reality. The mechanism for addressing the inadequacies defines the likely outcome of the project.CITATION:Lane, G and Skinner, E, 2013. Project delivery, in Proceedings MetPlant 2013 , pp 50-65 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 15, 2013

When the Middle Ages were drawing to an end and Rotterdam, then still a small fishing port, began to extend its trade relations to continental ports handling general cargo, a development was initiated which was to give Rotterdam a predominant position in maritime traffic in Europe 300 years later. Despite many setbacks and the silting up of the mouths of the Rhine and Meuse rivers, the port experienced a new rise towards the end of the previous century after the dunes at Hook of Holland had been cut through as planned by Dutch engineer Caland. With its favourable location in the mouth area of the Rhine and a hinterland opened up by waterways it was merely a matter of the city council pursuing a far-sighted policy to make the best of these natural advantages. Thus new port installations were erected on the southern bank of the Meuse in the first quarter of this century, which were expanded in the course of the subsequent decades. The New Waterway became the main artery for industrial enterprises, in particular oil industries, which sprang up along its banks (picture 1). Whereas 30 years ago Rotterdam's Waalhaven, with its new jettees for general cargo and mainly bulk goods, was still one of the most modern terminals, the growing size of ships, which increased by leaps and bounds for tankers, in particular towards the end of the 'fifties, necessitated much larger harbour facilities built further and further towards deep water.

Jan 1, 1972

By A A. Puhalovich

The placement of chemically reactive mine wastes in mined out pits (æpit backfillingÆ) is increasingly seen by industry as best practice to manage long-term environmental risks. This is primarily because leaving wastes at the surface, or allowing pit lakes to develop, can result in unacceptable environmental risks. The costs and practicality of mitigation options such as installation of covers and community expectations of landform re-construction to achieve a post-mining, beneficial land use are also important drivers for pit backfilling.Pit backfilling can significantly reduce the post-closure risks to receiving environments, by minimising seepage due to tailings, acid/solute generating waste rock and/or salts derived from mine water treatment. Environmental risks are often highest where wastes are located close to the ground surface, due to relatively high rates of groundwater movement in near-surface aquifers and/or failure of engineered structures due to rainfall/erosion processes. Placement of these wastes at significant depth in pits, subject to mine planning and environmental assessment, can address these risks by increasing the path length and resistance from contaminant source to receptor, thus enhancing encapsulation of the wastes.Pit backfilling can also be cost-effective for mining companies. This is because existing pits are developed on the basis of an economic return and thus provide an opportunity to store tailings in an existing æengineered structureÆ at little or no incremental cost when compared with building and operating a new tailings or waste rock storage facility, often encompassing complex cover/drainage management systems. Key issues for consideration are mine planning, the long-term geotechnical behaviour of the wastes (eg consolidation of tailings which leads to release of acid/solutes in tailings porewaters), seepage rates to downstream receptors/agreed end point criteria and potential resource sterilisation. Internationally, mine closure guidelines are also showing a greater recognition of pit backfilling in the context of allowing mining companies to meet agreed post-closure land use objectives.In summary, pit backfilling can enhance the encapsulation of chemically reactive mine wastes and environment protection, and with agreement on post-closure beneficial land use, can also help to deliver post-closure benefits to communities.CITATION:Puhalovich, A A, 2012. Pit backfilling - Managing chemically reactive mine wastes and achieving rehabilitation objectives, in Proceedings Life-of-Mine 2012 , pp 279-286 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 10, 2012

By M. Nzotta, L. Teng, S. Seetharaman, A. Chychko

The objective of the present work is to optimize the EAF practice towards Mo alloying with maximum retention of the alloying element in molten steel. During Mo additions, there were significant losses to the dust (ca 7 %) and to the slag phase (ca 2 %), which had serious economic and environmental impacts. The present work was aimed at designing a suitable Mo precursor, an alternative to ferromolybdenum, as the Mo source and a process sequence for optimal retention of Mo in the steel. The design of the Mo source should be in such a way that low cost raw materials such as MoO3 concentrate have to be used, stable during the addition so that the vapour phase losses as MoO3 are kept minimum and the dissolution of the precursor in molten steel is easy without loss to the slag phase. After thermodynamic considerations, three precursor materials were chosen, viz. CaMoO4, MgMoO4 and Fe2MoO4. Synthetic mixtures of the component oxides in each case as well as pure MoO3 were investigated by TGA and high temperature XRD in order to find out the temperature range in which the compounds were formed. Experiments were also conducted separately where the evolved gases were analyzed. All the three precursors were formed easily and the loss of MoO3 to the vapour phase was found to be extremely low. Addition of the oxide mixtures to steel melts were conducted in three different scales, viz. 16 g, 500 g and 3 tons (induction furnacle). Initial experiments showed that MgMoO4 was not suitable since the Mo yield in the steel melt was not very high. Fe2MoO4 was found to perform best, the Mo yield reaching 98 %. Further trials were conducted in the 65 ton Electric Arc Furnace (EAF) at Uddeholms AB in Hagfors. Mixtures of mill scale, MoO3 and carbon in suitable proportions were added in a pre-determined addition sequence. It was found that, in the actual practice, the Mo yield reached 99 %, equaling in performance the addition of ferromolybdenum. The process offers an effective use of cheaper raw materials and very high Mo yield avoiding a pre-treatment step like the production of ferromolybdenum. This is both cost effective and environmentally advantageous. The process has been patented and is being currently implemented at Uddeholms AB, Hagfors, Sweden.

Jan 1, 2011

By S. K. Jena, S. Singh, B. Das, N. Dhawan, P. K. Misra, D. S. Rao

"Iron as an impurity in pyrophyllite ore affects pottery, ceramics and tile quality. We carried out the characterization of a pyrophyllite ore sample using optical microscopy, scanning electron microscopy, X-ray diffraction, Fourier transform infrared spectroscopy (FTIR), thermogravimetric analysis (TGA) and wet chemical analysis. The study revealed the presence of 2.15% iron oxide (Fe2O3 ) along with silica and alumina as the major components, which constituted around 90% of the total sample. We carried out the removal of iron from the pyrophyllite using magnetic separation and acid leaching and found that wet high-intensity magnetic separation (WHIMS) could remove 28% of iron values while oxalic acid leaching could remove up to 99.3%. The leaching kinetics of iron removal and the effects of pulp density, acid concentration, leaching time and temperature were also investigated. The optimum conditions for leaching were found to be leaching time of 60 minutes, temperature of 90°C, pulp density of 5%, oxalic acid concentration of 0.3 M, and particle size below 100 microns. Iron dissolution from the pyrophyllite sample can be best described by the diffusion of ions through the product layer of constant-size spherical particles. The breakdown of the pyrophyllite structure was evident from the FTIR and TGA studies. The activation energy of the leaching process over the temperature range was calculated to be 43.47 kJ/mol. IntroductionPyrophyllite is a naturally occurring phyllosilicate mineral composed of aluminum silicate hydroxide [Al2Si4O10(OH)2] as the major component and often associated with other minerals such as kaolinite, mica, quartz, rutile and epidote. It occurs in fine-grained sericite schist that is quarried as pyrophyllite (Pandit et al., 2008). In India, pyrophyllite occurs mainly in Madhya Pradesh state, Uttar Pradesh state and the Keonjhar district of Odisha state (Harpavat, 1995). Pure pyrophyllite is a highly desirable material for various applications due to its unique properties. It exhibits high refractive behavior, low thermalelectrical conductivity, low expansion coefficient, low hot-load deformation, low bulk density and high corrosion resistance to molten metal and basic slag (Rieger, 1997). It is used in the ceramics, refractory, foundry, electrical-insulation, sanitary-ware, glassmaking, pesticide and rubber fields. Compared with clay, it has lower thermal conductivity and coefficient of expansion, making it more suitable for refractory applications (Harben, 2002; The Minerals Yearbook, 2011)."

Jan 1, 2015

"Metallurgical process development in greenfield projects must define the comminution flowsheet and establish the equipment sizes that are required to deliver the project design criteria. These are critical steps that heavily influence the financial performance of the project, requiring a detailed understanding of the ore properties of the project resources and the best practice methods to size comminution machines. This paper reviews current practices for determining the ore characteristics relevant to comminution circuit design. Particular attention is given to the design of sampling and testing programs, and the methods commonly applied to size crushers and mills for conventional size reduction applications in mineral processing. IntroductionMetallurgical process development in greenfield projects requires the selection of a comminution flow sheet and must establish the equipment sizes required to deliver the project design criteria. This necessitates a detailed understanding of the ore properties that define the project resource. For brownfield projects the same objectives may apply for plant expansions, or the project objective may be simply to define the capacity and operating cost of an existing facility for a new ore source. The process development phase in either instance will typically comprise the following key activities:• Geometallurgical classification and sampling of the ore deposit or deposits.• Determination of comminution energy requirements.• Benchmarking similar projects.• Identification of flow sheet options.• Equipment selection and sizing.This paper is concerned specifically with process development testing and will as such focus on the first two activities on the list: sampling and testing of the deposit. The discussion will focus on testwork programs to support the design of comminution circuits commonly in use in mineral processing today, specifically crusher, tumbling mill and stirred mill technologies.It is important that any testwork program is designed giving consideration to the types of flow sheets that are viable for the project, and the data that will need to be generated to support the final equipment sizing. Designing a testwork program to only provide information about one particular type of flow sheet will not allow sufficient information to be generated for a rigorous trade-off analysis of options, and may therefore preclude financial optimization of the project. Testing programs should be designed to collect sufficient data to ensure that any backtracking at a later stage to fill in missing information is minimal."

Jan 1, 2016

By S. S. Klima, J. A. Organiscak

Researchers at the National Institute for Occupational Safety and Health (NIOSH) performed laboratory testing to examine and improve airborne dust capture and methane removal performance of flooded-bed scrubbers on continuous mining machines using a blowing face ventilation configuration to reduce respirable dust exposure to shuttle car operators. Several testing conditions were examined to minimize shuttle car operator dust exposure without adversely affecting continuous miner performance and production. Curtain-to-face setback distance of 9.1 m versus 15.2 m (30 ft versus 50 ft), face ventilation airflow of 226.5 m3/min versus 339.8 m3/min (8,000 ft3/min versus 12,000 ft3/min), and the use of side-body-mounted blocking sprays (on versus off) were the three primary parameters compared. Scrubber airflow was set and maintained at 198.2 m3/min (7,000 ft3/min) throughout testing. An examination of slab cuts versus box cuts and the location of the shuttle car (present behind the continuous miner versus not present in the entry) was also performed to demonstrate the effects of these constantly changing scenarios on scrubber performance. Results show that the best blowing face ventilation setup for minimizing shuttle car operator dust exposure for the tested parameters uses a 15.2-m (50-ft) curtain-to-face setback with 339.8 m3/min (12,000 ft3/min) face ventilation while operating the side-body-mounted blocking sprays. INTRODUCTION Coal mine worker overexposure to respirable coal dust continues to be a problem in U.S. underground coal mining operations, potentially leading to coal workers’ pneumoconiosis (black lung). The U.S. Mine Safety and Health Administration (MSHA) requires that respirable dust exposure to coal mine workers does not exceed a 1.5 mg/m3 average over the actual working shift (79 Fed. Reg. 24814 (2014)). Although using a blowing face ventilation configuration on continuous miner (CM) sections is beneficial for removing methane emissions from the mining face, it can be detrimental for dust control for shuttle car (or ram car) operators and any other workers on the return ventilation side of the CM when compared to exhausting ventilation. Even with machine-mounted scrubbers on CMs, shuttle car operators still have some of the higher dust exposures with blowing face ventilation systems (NIOSH 2011, Beck et al. 2015). Although scrubber discharge air is not completely clean, dust being blown around the scrubber inlet and therefore not being captured by the scrubber can further increase these dust exposures (Goodman 2000). Dust control problems are also caused by inadequate face ventilation quantities causing scrubber face airflow recirculation issues (Schultz et al. 2010). In order to minimize respirable dust exposures on blowing face ventilation sections, ventilation, water spray, and scrubber parameters were experimentally examined to improve operating ranges for both dust and gas control. Several blowing face ventilation factors were examined with respect to CM spray systems and scrubber operation. Curtain-to-face setback distance and side-body-mounted blocking sprays were previously examined in the laboratory for dust control on a full-scale CM model using a scrubber on a blowing face ventilation system (Goodman 2000). The blocking sprays produced larger dust reductions at the CM’s rear corners and return for a 6.1-m (20-ft) curtain setback distance as compared to a 12.2-m (40-ft) curtain setback distance with equal face and scrubber airflows of 169.9 m3/min (6,000 ft3/min). Because face ventilation airflow quantities, shuttle car operator locations, and tracer gas concentration gradients were not previously examined during these experiments, additional research was required to determine improved operating parameters for blowing face ventilation systems.

Jan 1, 2019

By H. G. Stephenson, D. Castendyk

Open pit anthracite mining at the Canmore Creek Mines, Canmore, Alberta, south of Banff National Park, resulted in several pit lakes by the early 1970’s. To enable future mining opportunities in the Rocky Mountains, operators decided to reclaim the site into a non-polluting landscape which would provide a memorable outdoor experience for residents and visitors alike, and would preserve wildlife habitat. Reclamation began with the stabilization and revegetation of spoil piles adjacent to Canmore Creek, and the backfilling and revegetation of Canmore Creek Mines No. 1 and 2 to create “Salamander Pond.” The company dammed the outlet of an open pit at Canmore Creek Mine No. 3 to raise the water level and create Quarry Lake. Reclamation concluded in 1980 when the Town of Canmore took ownership. Quarry Lake is now the centerpiece of 293-acre Quarry Lake Heritage Park, a popular destination for swimming, hiking, and fishing, visited by about 400 people on summer days. The Alberta Environment Conservation Authority described this as “The best mine reclamation in Western Canada.” Quarry Lake is an exceptional example of walk-away, pit lake closure, where the mining company created a valued post-mining asset and was released from future liabilities. INTRODUCTION One of the most important environmental issues associated with contemporary open pit mining is the closure of mine pit lakes, and the perpetual legacy these lakes can potentially generate. Ideally, successful pit lake closure includes the following: (i) the water quality within the pit lake, plus surface water and groundwater discharging from the pit lake, meets pre-mining, or other acceptable water quality conditions of the receiving environment, (ii) the termination of all active treatment processes, (iii) the return of the pit lake to the government for long-term care and maintenance, (iv) the release of the mining company from all future liabilities associated with the pit lake, (v) the return the environmental bond to the mining company, and (vi) the reuse of the pit lake for some post-closure, beneficial purpose. The achievement of these goals is known as “walk away, pit lake closure.” Peer-reviewed, publications of successful walk away, pit lake closure provide mining companies the ability to address public concerns regarding the safety of pit lakes which commonly occur during the permitting, operations, and closure phases of mining. To date, most published work on walk away, pit lake closure work has occurred in the coal mining sector in Germany and Australia (McCullough et al., 2009; Jones and McCullough, 2011; Kumar et al., 2013; Schultz et al., 2013). However, in North America, examples of walk away, pit lake closure are not widely discussed in the literature despite there being excellent examples from the coal mining sector in Alberta, Canada (e.g. Pleasure Island, Black Diamond, East Pit, Sphynx, Robb, Lovette, Siltstone, and Sterling), and from the iron mining sector in Minnesota in the United States (e.g. Portsmouth Lake, see MNDNR, 2018). Although seldom included in discussions of pit lakes from the metal and coal mining sectors, the aggregate and fabricated-stone mining sectors also produce pit lakes, and there are numerous examples throughout North America where these pit lakes have been successfully closed and transformed into post-mining resources. The difficulty of finding these good examples in the literature makes it challenging for companies to respond to public and regulatory questions regarding pit lake closure and reuse. As such, there is a pressing need document the motivations and approaches used in examples of successful, walk away, pit lake closure in North America so that the mining industry at large can learn from these examples and can better address public concerns.

Jan 1, 2019

By Roshan B. Bhappu

Stimulated by current $0.80 to $1.20 per pound price tag of copper, considerable interest and activity have been generated in the copper mining camps throughout the world, and especially in the Southwestern United States and Northern Mexico copper belt. Under these economically favorable conditions, many of the marginal oxide-copper deposits, as well as newly discovered prospects, are under scrutiny for possible exploitation. For both the known deposits and new prospects, the big question facing the mine operator is the selection of the best method for processing these economically attractive oxide-copper deposits. Several variations of the hydrometallurgical processes are currently available for treating oxide-copper ores which can be mined by the open-pit method. Basically, these processes involve preparing the ore for application of an acid solution, applying the solution to the ore, and recovery of the metal from the pregnant solution. Dump leaching, heap leaching, vat leaching, agitation leaching and, in more recent times, in situ leaching or solution mining in combination with well established cementation process or the newly developed solvent extraction-electrowinning system have all been employed or have been investigated for the recovery of copper from oxide ores. Of these processes, dump, heap and vat leaching have been successfully practiced for treating oxide-copper ores. Agitation leaching is a relatively newcomer and has been adopted from the experiences of uranium and gold processing. Finally, the inplace leaching technique is the latest process to be suggested for possible treatment of oxide-copper ores.

Jan 1, 1974

By Renato G. Bautista

"The chemistry of the separation of metals from each other and from the other undesirable components involve multiple simultaneous chemical reactions. In the case of leaching, the heterogeneous chemical reactions between the solid and the liquid phase may be limited to a few reactions. In the homogeneous liquid phase, the number of simultaneous chemical reactions could be double or triple that of the number of heterogeneous reactions. The rate of dissolution becomes very dependent on the changing concentrations of the significant chemical species taking part in the multiple reactions in both the heterogenous and homogeneous phase. The rate expression developed for the process must take into account the mechanism of chemical reactions in order to best represent the actual process. This paper also examine the role of mass transfer with chemical reaction in the develpoment of the rate equations for dissolution, liquid-liquid extraction, membrane separation, precipitation, and electrodeposition.IntroductionThe understanding and the elucidation of the mechanism of the rate phenomena for a specific chemical reaction system is of prime importance in the development of a successful chemical process. The development of successful emerging technologies for metals separation will depend on the accurate interpretation of the intrinsic chemical reaction kinetics and the rate of mass transfer of the desired metal species during its separation from the undesirable chemical species. This is the case whether the process application is for separation, concentration, and /or purification.The chemical processes that will be examined in this presentation will include leaching, solvent extraction, membrane separation, precipitation, and electrodeposition."

Jan 1, 1996

By Christopher Jon Preston, Troy Williams, Daniel Roy

There has been significant progress made with narrow vein gold mines in which a second option for crushing has begun to attract attention. To ensure the profitability of a mine that has based its economics on the absence of a crusher station, more than just best blasting practices are required. Using a blasting operation itself as a primary crusher would undoubtedly save enormous capital costs as well as eliminate some of the costs of crushing ore with the result that considerable savings might be realized. The challenge would be to define the conditions necessary to make this type of operation feasible by providing consistent blasting methods that result in a significant paradigm shift with respect to both blasting and current underground crushing methods. A break model has been developed and applied as a technique in designing blasting patterns based on a unit charge blasting geometry and stress reflection from free faces. There are four components to this model and these are 1) unit charge, 2) thermodynamic/energy, 3) stress reflection, and 4) radial break. Some important changes in timing between holes and rows based on rock properties and the percentage of burden moved were also used. On-site testing at a narrow vein gold mine in Quebec Canada, provided measured parameters for the model. No underground crusher was planned - relying only on the fragmentation profile generated by ring blasting. The specification for muck was that the fragmentation profile needed to be 100% < 400 mm (16 in) with no overbreak - with dilution less

Jan 1, 2019

By M. Mpofu, B. Maphalala, S. Khan, S. Ngobese, D. Roberts

This paper details an assessment of stemming practices at a South African opencast coal mine and their influence on ground vibration and air blast. Quantitative and qualitative analysis methods were used for the study. The parameters assessed for the quantitative analyses included stemming length, stemming material type, blast-hole depth, burden, and spacing. Pre-blast data from the mine was analysed to determine the deviation between actual and mine design stemming parameters. Mine design stemming length was also compared to the rule-of-thumb stemming lengths. Peak particle velocity (ground vibration) and air blast data from seismograph stations around the mine was analysed to determine the correlation between stemming length and excessive ground vibration and air blast. The qualitative analysis involved observations of blasting procedures at the mine to determine compliance with mine procedures. Some level of non-compliance of actual stemming parameters to design stemming parameters was found, which initially suggested that non-compliance may have contributed to excessive air blast. However, further analysis of the seismograph results indicated no direct correlation between stemming length and excessive air blast and ground vibration. Since stemming parameters are interrelated, it is crucial that all blasting procedures, including stemming, are executed in accordance with best practice, and recorded accurately. There is a need for digital systems for capturing on-bench blasting parameters, as the currently used manual data recording and reporting systems are inefficient and prone to error. Improvements to blast designs are possible through efficient and accurate data recording and reporting systems.

Jan 1, 2021

By K. Thuro, L. Wilfing, H. Kasling

"In mechanized tunneling, performance prediction is a crucial issue for estimating excavation costs and construction time of a tunnel project. The aim of this work is to compare the calculated parameters by penetration prediction models with the results of penetration tests during excavation by a Tunnel Boring Machine (TBM) at actual tunnel projects. This comparison helps to improve the existing prediction models (Gehring, 1995 and Rostami (Colorado School of Mines), 1997.) and to develop a new, so-called “Alpine Model” (research group ABROCK; Schneider et al., 2012).On-site penetration tests are a common tool to determine performance of a TBM in certain geological environments. During a test, the TBM is operated under defined conditions that allow comparing the results of different projects and machine types. There are two different types of penetration tests in use. The more common type, described by Bruland (1998a), is based on test drives at different thrust levels. The tests in this study are performed as start-stop-tests, meaning that the TBM starts from zero to the maximum possible thrust and reverse with slowly increasing respectively decreasing steps (Frenzel et al., 2012). To gain information about the interaction between TBM and rock mass, a detailed geological documentation including an analysis of the discontinuity pattern and laboratory testing for rock strength characterization are mandatory.Within the last year 16 penetration tests have been performed. The results are thrust- penetration curves (Figure 2) showing that up to a certain penetration rate only crushing occurs (subcritical penetration). From there the desired rock fragmentation process called chipping takes place (Frenzel et al., 2012). These two different fragmentation processes can be described best by a bilinear function as seen in Figure 3. Both, CSM- and Gehring-model, do not fit to the trend of the penetration rates determined during the excavation process as they are using potential or simple linear functions.Another parameter that has to be implemented into the new prediction model is the influence of discontinuities. In case of favorable orientation and spacing, discontinuities reduce the required thrust at same penetration rates. In contrast, unfavorable discontinuity systems can cause blocky faces and hence worse performance than predicted. Besides discontinuities, the toughness of rocks has to be taken into account when predicting the penetration as tough rocks need more energy to obtain complete failure than brittle rocks.Key advantage of these tests in comparison to existing are more data points in less time. By improving respectively extending the existing penetration prediction models with the stated parameters, the penetration prediction for TBMs can be more accurate and also wear calculation can benefit from this approach."

Jan 1, 2015

By J I Stewart

The Copper Canyon deposits, occur in proximity to the regionally distinctive formational boundary between the Staveley Formation and Marimo Shale. The mineralisation here is associated with breccias, veins, and alteration zones near the base of the Marimo Shale unit and in the underlying evaporitic dolomite-dololutite dominant Staveley Formation. The Marimo Shale unit, in its basal 60 metres, is highly carbonaceous and hosts ubiquitous but minor diagenetic (?) copper-pyrite, and on occasion, extensive albite-microcline alteration centres, pyritic (cobaltiferous)-quartz veins, and microcline-quartz_+plagioclase veins or veinlet arrays. The most significant mineralisation, such as at the Greenrnount prospect, consists of a shallowly dipping package of alteration zones overprinted by multiple generations of pink-orange microcline-quartz-chalcopyrite-pyrite-(arsenopyrite) and albite-calcite- phlogopite veins. Accessory mineralogies in most of these mineralisation styles consist of tourmaline, monazite, rutile, and apatite. Chalcopyrite (and secondary chalcocite) and gold are associated with each generation of alteration but are best developed in the later stages of veining. In the underlying Staveley Formation, which is a conformable transition between the Corella Formation limestones and the Marimo Shale, chalcopyrite (cobaltiferous)-pyrite mineralisation occurs as disseminations in pre- and syntectonic breccias, and with disaggregated silica pseudomorphs (after thinly bedded evaporitic sulphate beds) and associated dolomitised limestone-calcilutite. Although this mineralisation may be partly diagenetic in origin the 200 in thick evaporitic packages are also host to a locally widespread (2 to 4 km long and 200 to >500 metres wide) and pervasive &apos;redrock&apos; styled microcline-dolomite¦hematite alteration. Within this gross alteration package there also seems to be a consistent association with narrow (10 to 100 metres wide), linear (stratabound) semi-massive magnetite-hematite-microcline and quartz-chalcedony replacement zones. Like much of the nearby Corella Formation the Staveley Formation, under the Marimo Shale, has undergone dissolution, paleo-solution karsting (breccias) and tectonic brecciation such that the precursor lithological details have been progressively obliterated. The post-diagenetic to post-tectonic fluid pathway development and metasomatic alteration history of this important lithostructural package of rocks has been traced from preliminary studies only. The ability to do so is largely due to the absence of significant regional and contact metamorphic overprinting, and the local preservation of precursor sediment profiles. It is proposed that both the base of the Marimo Shale and parts of the underlying evaporitic-limestone-dololutite have acted as conformable fluid pathways for hydrothermal and (perhaps) diagenetically derived, Cu-Au-(¦Co ¦As) K-Na-Ti-REE bearing, fluids and brines.

Jan 1, 1994

By K. C. Corin, Z. G. Song, J. G. Wiese

In the present investigation flotation was used as a diagnostic to understand the effects of changes to the media during the grinding stage. The observations from flotation for different media are discussed in terms of particle size distribution, water recovery, solid recovery and valuable mineral recovery. The chemical parameters; pH, redox potential (Eh) and dissolved oxygen (DO) levels were measured. A Base Metal Sulphide (BMS) ore was ground under media of different shapes and compositions to determine the effect of different rod/ball material on the flotation performance of the ore. For a given grind (D80 = 75 µm), media shape had no effect on flotation performance of a BMS ore. Different grinding media had significant effect on slurry chemical properties and flotation performance. Mild steel rods (RMMS) produced the highest solids recovery, nickel and pyrrhotite (Po) recovery, with the highest ethylenediaminetetraacetic acid (EDTA) extractable iron and the lowest DO level. Chrome steel balls and stainless steel rods produced the lowest metal recovery, EDTA extractable iron and the highest DO level. A complementary study was conducted to investigate the effect of using different grinding media compositions, viz, ceramic, forged steel, 15% chrome steel, stainless steel and a mixture of stainless steel and 15% chrome steel (50:50) balls were investigated using a Platinum Group Mineral (PGM), UG2 ore. Grinding products of five different media produced the same solids and water recovery, but the grades and recoveries of Platinum Group Elements (PGE) and chromite were different. Grinding products of ceramic media showed the best flotation performance, with the highest PGE recovery and the lowest Cr2O3 recovery. While grinding products of stainless steel balls (SS) produced the lowest PGE recovery and highest Cr2O3 grade and recovery in the concentrates.

Jan 1, 2016

By E. E. Mason

THE Myrtle group of mining claims is situated on Barkerville mountain in the Cariboo mining district of British Columbia, west of the old mining town of Barkerville. An 1,800-foot adit, known as the Shamrock tunnel, was driven in 1935-36 from the Barkerville road and in the general direction of the Myrtle group. It was proposed to enlarge this adit to more convenient section, extend it 2,500 feet int0 Myrtle ground, and from there drive a main entry 3,000 feet along the length of the group. Several laterals were also to be driven off this main entry as part of the tunnel schedule. It was thus originally proposed to drive 5,500 feet of new adit and 1,800 feet of laterals, a total of 7,300 feet of new work. Work was commenced with a bulldozer on April 21st, 1941, cutting a bench on the hillside for a plant site. The power plant was ready for operation on August 8th. Slashing and re-timbering of the old adit rook longer than was expected. It was found that 57 per cent of the length required timbering, chiefly by fore-poling, and there was also trouble in several sections with swelling ground. First round in the new face was taken on September 17th, when a 2-shift daily schedule was commenced. Due to delays in delivery of electrical equipment, it was not possible to add the third shift until November 14th. From then on, a 3-shift daily schedule was maintained with the exception of holidays and fortnightly change Sundays until August 1st, 1942, when shortage of labour caused the dropping of a shift. All work was suspended on August 14th. The ground through which the tunnel passed was such that the driving was often interrupted by the need of protective timber at the face. Further, crews were continuously employed timbering behind the face, and in maintenance of timbered sections in swelling ground. Roughly 980 feet of the new workings are now timbered, about 21 per cent of the total.

Jan 1, 1942

By S Graskoski

The allocation of boggers and trucks at Telfer’s sublevel cave mine is currently assigned by manual techniques, with decision-making primarily dependant on the judgement of the operators and mine control. Field observations indicate that allocation decisions could be improved by establishing a systematic approach to dispatch. An investigation into the current state of technology determined that no currently available solution would provide a tool to make optimal decisions for the Telfer dispatch challenge.In an underground sublevel cave (SLC) environment, the destination decision is fixed and the source is variable, so open pit and block cave solutions are not relevant. Newcrest underwent a development venture with an external vendor to develop a novel production dispatch algorithm and user-friendly software program to make allocation decisions in a systematic way with consideration given for the true complexity of the operation.This new production software, nicknamed Dingo (for Dispatch ‘n’ Go), is a smart way to assist forepersons and mine control operators in their day-to-day roles of planning shifts and allocating trucks and loaders to optimal locations during shifts.CITATION:S Graskoski, 2017. Dingo – an optimisation tool to solve complex short interval control at Telfer underground mine, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 261–266 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 16, 2017

By D. G. Sinclair

IN surveying the changes that have taken place in the mining industry of Ontario in the past ten years, the striking growth of the gold mining industry first attracts attention. In 1925 there were only twelve regularly producing gold mines in the Province, six of which were located in the Porcupine area, four at Kirkland Lake, and one each at Larder Lake and Boston Creek. In contrast with this, the year that has just closed (1935) has seen production on a steady basis from forty mines, with what may be termed spasmodic production from twelve other properties. This so-called spasmodic production is, however, responsible for the recovery of gold to an estimated value of between $75,000 and $100,000 during the year. Further illustrating the expansion in the gold mining group is the activity in the exploration and development of properties. In 1925, mining operations were conducted at 29 properties which had not reached production, while in 1935 work of this nature was carried on at 87 such properties. It is of interest to note that of the six properties listed as producers in Porcupine ten years ago, five are still producing at rates considerably in excess of the tonnages handled at that time. In fact, the estimated tonnage treated at these mines in 1935 exceeds the 1925 tonnage by approximately half a million tons. In Kirkland Lake, the figures are still more striking. The four mines in production in that camp in 1925 milled 346,584 tons, producing bullion to the value of $5,132,128, while in 1935 these same mines treated approximately 1,643,000 tons, with a recovery of roughly $29,300,000. While Porcupine and Kirkland Lake remain the great gold-producing camps of the Province, the growing importance of the newer areas is readily appreciated when we consider that in 1935 the mills of these areas treated over 1,000,000 tons of ore for a production in excess of $7,500,000. Pro-duction from these newer districts first became of importance in 1930 with the commencement of recovery at the Howey mine, in Red Lake, and the Minto, in Michipicoten. Since 1930, eighteen new gold mills have gone into production in the Red Lake, Little Long Lac-Beardmore, Pickle Crow, Sturgeon Lake, Michipicoten, Sudbury, and Matachewan areas. In respect to silver mining, the period under review has seen a considerable slackening of activity, due to a generally unstable and depreciated price for the product and to the exhaustion of ore reserves in the mines.

Jan 1, 1936

By G. R. Addie

While the cost of transporting raw phosphate matrix from a pit to a processing plant via slurry pipelines is more efficient than it could be by rail cars, conventional trucks or belt conveyors, it still amounts to many millions of dollars, and in many areas there still exists much room for improvement. By taking advantage of pipeline test data on different slurries, it is possible to reduce costs by optimizing pipeline diameters, operating velocities and slurry concentrations. Better wearing, more efficient pump designs are now available which, if applied correctly, can lead to significant savings. The authors discuss what benefits are possible in relation to the electrical operating and wear maintenance costs and go into some of the factors associated with taking advantage of these benefits, identifying what they see as being practically possible.

Jan 1, 1998

By Kaustubh Joshi, Pradip, Beena Rai, Kesri L. Menaria, Venugopal Tammishetti

"Beneficiation of alumina rich iron ore slimes is a major challenge for the Indian iron ore industry. High alumina is deleterious for the performance of both blast furnace as well as sintering plants. Slimes generally contain 5 - 10% alumina, and hence despite assaying as high as 60% Fe, those are dumped in tailing ponds. The slimes are typically less than 37 microns. Selective flocculation of iron ore slimes using highly selective reagents is thus an option worth exploring. Our research is directed at employing state-of-the-art molecular modeling techniques for the selection/design of selective reagents for this separation. We have earlier reported our results on starch as a selective flocculant for iron ore slimes. In this paper, we present our results on the selective dispersion-flocculation of natural iron ore slimes using guar gum flocculant. With careful selection of various process parameters such as flocculant dosage, dispersant dosage, pH and settling time, we were able to obtain a concentrate assaying >66% Fe and <3.5% Al2O3 with an yield of 65 to 68% (recovery of iron values more than 70%) from a natural iron ore slime sample assaying 58% Fe and 7% Al2O3."

Jan 1, 2014