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By Malcolm Scoble, Jean-Marc Robert, Luciano Piciacchia

"An approach to the in situ testing of mine backfills is outlined. It has been developed to provide a measure of the effective strength and stiffness of such materials after placement in stopes. This paper is based upon recent field studies undertaken at five Quebec mines. Details of the selected instrumentation, supporting equipment and testing procedures are reviewed. The mechanical behaviour of the backfills is reported and related to their physical properties. This approach represents an economical, simple and reliable means to assess the mechanical characteristics of backfill materials as an aid to both mine design and placement quality control. IntroductionDevelopments in backfill and mining method technology are creating a pressing need to develop effective techniques for characterizing the physical and mechanical properties of backfill in situ, rather than relying on traditional laboratory testing. The development of bulk mining methods is creating the requirement to more fully understand the nature of exposed fill faces to design for stability in pillar recovery operations. Other backfill technological developments currently receiving widespread attention are high density placement and the use of alternative additives to Portland cement. In Canada, U.S.A. and South Africa attention is also being placed on employing backfill to alleviate rockburst hazards(1,2.3l. In situ testing offers the potential to measure the geomechanical properties of backfill once placed in stopes , in order to assess its actual role in pillar recovery and ground support. Similarly, it offers the ability to provide some quality control during backfill placement. This role is seen as becoming increasingly important with the growing recognition of the need to pay greater attention to the mode of placement, which governs segregation and variation in the properties actually achieved by the backfill after placement."

Jan 1, 1987

A large lab apparatus was designed using a scale-down methodology based on the design legacy of the current FLSmidth commercial forced air flotation machines. This new apparatus, along with a new test methodology, was used to evaluate new rapid prototype concepts. The apparatus was designed to monitor bubble size, power, pumping capacity, Jg, velocity profiles, froth character, and tip speed. The methodology developed follows a sequence that goes from hydrodynamic tests and solids suspension tests to a flotation kinetic test. This methodology has been demonstrated to be a good tool to determine the effect of machine design changes on flotation.

Jan 1, 2012

By Leonardo Flores

"Minera Escondida operates two open pits on a porphyry copper district, southeast of Antofagasta, Chile. During 2005, ore will be delivered exclusively from one pit to two concentrators, while from 2006 both pits will provide ore to these plants. Annual testing programmes on drillhole samples, have allowed for the construction of hardness block models since 2000. These models have delivered two main hardness descriptors, the Bond work index and the SAG power index. These parameters and plant operational specifications have been used as input information to forecast, the expected throughput and ball mill product size (P80) for each block of ore. Two forecasting techniques have been used, an in house model, strongly dependant of bond work index and MinnovEx´s CEET2, dependant of Bond work index and Sag power index and other parameters. Model performance has been evaluated from reconciliating on a monthly basis, the forecasted throughputs against a recalculated plant throughput which accounts for losses in tonnage not caused by ore hardness but instead by unscheduled plant operation problems. The model performance for 2003 was ± 2% yearly and ± 5% quarterly, well within the expected ± 3% and ± 10%. INTRODUCTIONThe Escondida porphyry copper district is located 140 km SE of Antofagasta, Chile at 3150 m.a.s.l, at 24°16’ South latitude and 69°04’ West longitude (figure 1)."

Jan 1, 2005

By Gunhild Hennum, Tewodros Tefera, Gisle Håland, Annette Jahr

"Abstract The Gudrun field is located west of Stavanger in the North Sea at water depth of 109 m. A jacket substructure was selected to support the topside and was installed in 2011. The jacket shall be able to support 16 of 0.762 m (30 in) diameter conductors and is equipped with a drilling deck to allow for pre-drilling by a jack-up rig through the jacket. In general the soil conditions below mudline comprise of a thin top layer of very loose to dense sand overlaying several layers of sandy clay, ranging in stiffness from firm to hard. Below some 30 m depth the soil consists of layers of medium dense to dense silty sand overlaying very hard sandy clay containing thick beds of sand to 100 m depth or more below mudline. The jacket is secured to the seabed by twelve 2.438 m (96 in) diameter steel pipe piles, three at each pile cluster located at the four legs of the jacket at sea floor. Different design methods were used to determine the pile penetration depth. Pile drivability was evaluated using best estimate and upper bound soil parameters. The pile driving and blow count record show that the target penetration was reached as predicted. This paper discusses the different methods used to determine the pile capacity, target penetration depth and the drivability analyses made for the Gudrun jacket piles. The soil resistance during pile driving is back-calculated from the pile installation data and compared with the prediction made.IntroductionThe Gudrun platform is located west of Stavanger in the North Sea at water depth of 109 m. A jacket substructure was selected for the Gudrun platform. The general Gudrun platform jacket configuration is to a large extent dictated by the topside layout, the requirements for support points enclosing the well slots and access for the jack-up rig. The Gudrun platform jacket is a traditional 4 legged structure with a vertical side towards the jack-up rig. At the seabed the jacket is quadratic with a base dimension of 40 by 40 metres. Figure 1 shows the Gudrun jacket during load-out from the fabrication yard to the transport barge. At the site the jacket was lifted off the barge, upended and placed on the seabed by the installation vessel. The jacket is secured to the seabed by twelve 96"" (2.438 m) diameter steel pipe piles, three at each pile cluster located at the four legs of the jacket at sea floor. The target penetration of the piles is 47 m."

Jan 1, 2014

By D Craw, P Upton

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. The Otago Schist belt of southern New Zealand hosts widespread orogenic gold deposits. Erosion of these orogenic gold deposits has led to the formation of numerous placer deposits distributed across and beyond the host schist belt. Numerical modelling of crustal scale deformation and geomorphic evolution has defined the key tectonic features that have controlled the severance of detrital gold deposits from their sources. Tectonic activity over the past 20 million years, during development of the modern Pacific-Australian plate boundary, has severed links between some placers and their sources. Detrital gold was transported south from the Otago Schist source region in two different river systems in the Miocene. Mesozoic orogenic gold deposits shed gold into rivers that flowed to the south. These rivers were terminated by the rise of north-west-trending basement ridges at the boundary between weak schist crust and strong greywacke crust, and the ridges were controlled by pre-existing faults at that rheological boundary. In contrast, rising mountains close to the developing plate boundary in the Miocene were the locus of new orogenic gold deposits hosted in extensional faults. These gold deposits were rapidly exhumed and contributed detrital gold to orogen-parallel Miocene rivers that flowed to the south-west. Pleistocene deformation on mainly north-west and north-east orientations, and associated river capture, has severed the links between these Miocene orogenic deposits and downstream detrital gold.CITATION:Craw, D and Upton, P, 2015. Tectonic severance of links between placer gold and its sources, southern New Zealand , in Proceedings PACRIM 2015 Congress, pp 541–544 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By Pr-Anders Persson

This chapter deals with the concepts of shock waves and detonation waves together, because a detonation wave is really a shock wave, supported by the explosive reaction that the shock wave ignites and propagates as it travels through the explosive material. In the c.ootext of rock blasting, furthermore, there is justification for treating shock waves and detonation waves together, because the mechanical effect of the detonation in an explosive in a drillhole in rock is propagated and its effects are spread throughout the surrounding rock by means of a shock wave. The overall performance of sn explosive in rock blasting or any other application is of course dependent on how much thermal energy is released in the chemical reaction that occurs when the material detonates, and on how much useful mechanical work the reaction products can do as they expand. A major portion of this chapter deals with the prediction and measurement of explosive performance, a subject which is now still in a process of rapid development.

Jan 1, 1995

This paper describes a proposal to reopen the Sylvia gold mine for underground exploration and development for an initial production of 10 000 tpa. A combination of trackless mining equipment and railtrack would be utilised for development with ore extraction being carried out by undercut and fill methods. The use of the smallest LHD units for mucking, and rail track for the main haulage together with airleg drilling will form the basis for development and stoping. The proposal has been subjected to an environmental impact assessment report to evaluate the impacts of the proposal on the natural and social environment in the area.

Jan 1, 1987

By E. Concha

The concept of an Automated Continuous Explosive Excavation system (ACEE) for the underground mining applications of development drifting and room and pillar mining has been evaluated in comparison with conventional drill, blast and muck operations. The role and possible extent of application as this method was identified. The evaluation examined each element of existing and of proposed ACEE design concepts in a defined mine environment. Step by step procedures were developed to quantify the evaluation process where this appeared feasible. One conclusion drawn was that a system approach to ACEE design is not only feasible but attainable within the state of the art. Several preliminary design concepts were formulated and examined in order to determine what further research or development might be needed to realize them.

Jan 1, 1982

By R. V. Taborelli, P. B. Bucky

This investigation was carried on by means of models and the application of the principles of similitude to determine the effects of props, props and cribs and sand filling in longwall mining. The geologic structure and rate of face retreat were kept constant and the only change introduced was in the method of support. It is not felt necessary to repeat a description of apparatus, or to explain the fact that a scalar model made in all parts of the same material as the prototype will behave similarly to the prototype when rotated at speeds such that the centrifugal force is as many times the gravitational force as the prototype dimension is a similar linear model dimension. The reader is referred to earlier papers1,2 for further information regarding these statements. All models were made of sandstone and were as in Fig. 6. They are composed of a 3/16-in. ore body bed A, overlain by a 3/16-in. bed B, a ?-in. bed C and two ?-in. beds D and E. The model was rotated at + 2350 r.p.m. equivalent to a model ratio of 1268. It was worked back ¼ in. at a time before being replaced in the centrifuge and again rotated. Table 1 gives model and prototype dimensions occurring in this discussion, the prototype dimensions being 1268 times the model dimensions. Figs. 1, 2, 3 and 4 were taken while the model was rotating at 2350 r.p.m. One was taken for each face retreat of 26.4 ft., but for obvious reasons all are not presented. Fig. 1 (A-E) shows the behavior of this type of structure with face retreats of 26.4 ft. and no artificial support. Fig. 2(A-E) shows the behavior of this type of structure where breaker and face props were used. The props were exceptionally strong, being of steel ? in. square in cross section, and capped with 1/16 in. ± of soft wood and paper, to allow for roof deflection. The face props were

Jan 1, 1940

By P. B. Bucky, R. V. Taborelli

This investigation was carried on by means of models and the application of the principles of similitude to determine the effects of props, props and cribs and sand filling in longwall mining. The geologic structure and rate of face retreat were kept constant and the only change introduced was in the method of support. It is not felt necessary to repeat a description of apparatus, or to explain the fact that a scalar model made in all parts of the same material as the prototype will behave similarly to the prototype when rotated at speeds such that the centrifugal force is as many times the gravitational force as the prototype dimension is a similar linear model dimension. The reader is referred to earlier papers1,2 for further information regarding these statements. All models were made of sandstone and were as in Fig. 6. They are composed of a 3/16-in. ore body bed A, overlain by a 3/16-in. bed B, a ?-in. bed C and two ?-in. beds D and E. The model was rotated at + 2350 r.p.m. equivalent to a model ratio of 1268. It was worked back ¼ in. at a time before being replaced in the centrifuge and again rotated. Table 1 gives model and prototype dimensions occurring in this discussion, the prototype dimensions being 1268 times the model dimensions. Figs. 1, 2, 3 and 4 were taken while the model was rotating at 2350 r.p.m. One was taken for each face retreat of 26.4 ft., but for obvious reasons all are not presented. Fig. 1 (A-E) shows the behavior of this type of structure with face retreats of 26.4 ft. and no artificial support. Fig. 2(A-E) shows the behavior of this type of structure where breaker and face props were used. The props were exceptionally strong, being of steel ? in. square in cross section, and capped with 1/16 in. ± of soft wood and paper, to allow for roof deflection. The face props were

Jan 1, 1940

By R. Ciccu, P. P. Manca, A. Bortolussi, G. Massacci

Tool performance and the unit cost of cutting depend on a number of operational variables (peripheral velocity, pull-back force, water injection rate, angles of the sections of wire outside the rock) as well as on the geometric features of the cut itself (area and shape). Cutting rate and wire productivity are also influenced by the type and number of diamond beads per unit length of wire and by the characteristics of the rock (granite or marble). A computer program aimed at handling a problem of this complexity has been developed and tested with the use of available field and laboratory data and has been found acceptably reliable. Discrepancies between current industrial data and some of the results suggest future lines of research, while the simulation should also be extended to include different types of rock and tool and a variety of bench face geometries

Aug 1, 1994

By G McMillan, P Mulvey

"Geochemical characterisation is done based upon the geology and metallurgical batch testing prior to commencement of the operation of the plant.  Tailings repository design is based upon geochemical assessment of the waste material, as well as the geomechanical issues associated with storing the waste. The resultant conceptual and detailed design is incorporated into the environmental impact statement (EIS), upon which mine approval is based. The entire environmental monitoring and harm mitigation program, which is usually a condition of the Mine Lease approval, is based upon the applicability of the original characterisation.Tacit to this process is that there is no change in the geochemistry of the tailings, i.e. no change in the implications for acid & metalliferous or saline/neutral mine drainage (AMD/SNMD) generation. But any changes in milling and processing can have a substantive impact on the geochemical behaviour of the tailings metallurgy, and potentially render an operation non-compliant within the conditions of the mine license.Environmental and statutory approval implications for continuous improvements of metallurgy need to be understood because unintended consequences could cost more than the benefits of the improvement, and in some cases cost more than all the profit made from the mine.  Examples uncovered during audit and consultant practice will be presented together with a process to ensure implications for adverse environmental harm and license conditions are fully considered.*This paper is only available for individual download. It is not included in the conference proceedings*CITATION:Mulvey, P, 2013. Continuous improvement - how changes in metallurgy can have a substantive impact on environmental impact and closure - lessons learnt from environmental auditing, presented to MetPlant 2013 Conference (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jul 15, 2013

Eco-Efficient Gravity Separation Solutions in India Sustainable development and social responsibility issues are a principal aspect of mineral processing operations around the world. With the need for more environmentally effective technologies, some gravity separation systems have provided eco-efficient mineral processing solutions in places where it is not otherwise viable because of lack of fresh water or for other environmental reasons. Advanced spiral separation technologies developed in Australia offer a cost effective and environmentally viable way to process minerals. In particular, we review recent case studies of spiral plants commissioned in India where environmentally sustainable solutions to design problems included the use of spirals operated with sea water and the use of high capacity (HC) spirals to provide low consumption of energy and materials and a reduced plant footprint. This paper evaluates the performance of various spiral plants in India and how critical environmental issues which affect the overall viability of wet concentrating processes where resolved.

Sep 13, 2010

By George K. Burke

Jet grouting was first introduced into North America in the late 1970's, soon after the conception of the system in Japan. However, it was not until the mid-1980's that the technique began to gain acceptance through the renewed interest and efforts of specialist U.S. geotechnical contractors. Since this time, innovations by both American users and other users worldwide have further advanced the technology, with a corresponding increase in the range of practical applications. This paper discusses the evolution of jet grouting in North America, from simple single-fluid methods to more complex triple-fluid methods, to the energy-intensive double-fluid and SuperJet methods. Case histories, illustrating a variety of jet grouting methods and applications, are cited to support the newest United States applications.

Jan 1, 2002

By D N. B Skinner, R L. Brathwaite, M S. Rattenbury

The Institute of Geological and Nuclear Sciences Ltd has carried out a mineral resource assessment of the Coromandel region, one of the most mineral-rich regions of New Zealand. The assessment focussed on metallic minerals and clays, because these are the main commodities that have potential for increased production and export from the region (and from New Zealand).   Three steps were used in the assessment: 1) development of mineral deposit models, 2) comparison of the models with regional geology, structure, geochemistry and past mineral production to estimate the number of undiscovered mineral deposits, and 3) application of grade-tonnage models to determine the quantity and value of the known and undiscovered resources. In step 1, descriptive and grade-tonnage mineral deposit models were developed for 11 different types of mineral deposits: ò     Low-sulphidation andesite-hosted epithermal Au-Ag ò     Low-sulphidation rhyolite-hosted epithermal Au-Ag ò     High-sulphidation quartz-alunite epithermal Au ò     Porphyry Cu-Au-Mo ò     Andesite-hosted polymetallic veins adjacent to mineralised porphyry stocks ò     Zn-Pb-Ag skarn ò     Cu-Ag skarn ò     Hot spring Hg ò     Shoreline placer Ti ò     Rhyolite-hosted halloysite clay ò     Volcanic-related kaolinite clay   Step 2 was aided by a geographic information system (GIS) assembled from a number of digital spatial data sets including: a digital elevation model, an aeromagnetic map (from open file mining company data), a geological map (from the GNS QMAP project), location of areas of hydrothermal alteration, locations of mineral deposits (from the GERM database), stream sediment geochemical analyses (from the REGCHEM database), and a lineament map derived from satellite remote sensing data and the digital elevation model.   Based on May 2001 commodity prices, potential resources estimated for the 11 mineral deposit and grade-tonnage models total NZ$22 806 million for the metallic minerals, NZ$1000 million for halloysite clay deposits and NZ$57 million for kaolinite clay deposits.

Jan 1, 2001

By A Croll

A panel cave using LHDs for ore extraction came into production at Premier in 1990 at a depth of 630 metres below surface on the western side of the mine in the BA5 mining block. A post-undercut mining sequence was used with all development on both the undercut and production levels completed prior to the undercut being advanced over the pre-developed drawbells. Support and rock mass damage on the production level was extensive as the undercut moved overhead. Undercutting and coarse fragmentation created problems. The mining sequence was changed to an advance undercut with the minimum of development and support undertaken on the production level prior to the undercut moving overhead. Only once the abutment stresses had stabilised were drawbells developed and final support installed. The height of the undercut drill design was decreased from 18 to 13 metres. The changes resulted in markedly less rock mass and support damage on the production level. Lessons learned in the BA5 mining block on the 630 level were incorporated into planning and scheduling the BB1E mining block which came into production on the 732 metre level on the eastern side of the mine in 1996. An advance undercut mining sequence was planned, undercut height was minimised and undercutting parameters were optimised in terms of the BA5 experience. Numerous problems still occurred. As the undercut area increased in size severe spalling from the cave back resulted in vertical propagation of the cave to a height of 30 metres instead of the planned 13 metres. The resultant coarse fragmentation created secondary blasting problems and an excess of undercut ore that had to be loaded and trammed, slowed the rate of undercut advance. The planned mining sequence was compromised and considerable support and rock mass damage resulted on the production level below. Once these problems had been overcome the undercut advanced at the planned rate and continuous caving initiated at the predicted hydraulic radius. At a later stage the unexpected ingress of water hampered undercut drilling and slowed the rate of undercut advance in a section of the cave. Leads and lags between adjacent tunnels and the overall face shape could not be developed as planned. The rate of natural horizontal cave propagation exceeded the rate of undercut advance and three undercut drilling tunnels were extensively damaged. More tunnel development on the production level prior to undercutting than was planned resulted. The slow advance of the undercut and high abutment stresses caused extensive damage to installed support and the rock mass on the production level, 15 metres below. Experience in the BB1E mining block on the 732 metre level has taught that an advance undercut mining sequence reduces damage on the production level markedly. Extremely tight control on the rate of undercut advance, leads and lags between adjacent tunnels and overall undercut face shape as well as extraction ratio, tunnel development and support sequence on the production level below is required. Premier Mine plans to implement a block cave at a depth in excess of 1000 metres within the next decade. All the experience gained on current mining levels together with data collection and numerical modeling is being used to plan the proposed mining operation. The geotechnical challenge of mining at this depth in weak rock has been identified as the greatest risk to the project. Challenges centre around adopting the correct mining sequence, limiting the extent of rock mass damage associated with undercutting and effective support. The hydraulic radius needed to initiate caving, fragmentation, drawpoint spacing, secondary blasting and the possibility of seismicity are geotechnical concerns that are being addressed. Results to-date suggest that a pre-undercut mining sequence with all development and support on the production level carried out only after the undercut abutment stresses have passed overhead will have to be implemented.

Jan 1, 2000

By Henry S. McDevitt

External Shock Mitigation (ESM) is being studied as a possible means of retrofitting existing buried structures to provide additional protection from ground shock. As part of this study, two tests were conducted on four ESM models and a control structure. The test articles were placed at a carrnon distance and subjected to the ground motions created by a buried explosive charge consisting of 2,ODO pounds of nitranethane (NM). Adequate instrumentation data were obtained to assess the response of the four ESM models. The assessment relied heavily on direct comparison of horizontal motions measured in the models, with similar measurements in the control structure. From the tests, it was observed that peak accelerations were reduced by a factor of 5 by the ESM. However, the particle velocities and displacements were not significantly affected by theEIS.

Jan 1, 1997

The objective of this study was two-fold: 1. To perform a world literature search pertinent to the analysis, selection and design of hoisting systems and their components and to codify the search into a literature compendium. The search resulted in a technical data file of over 1500 hard copy items. The compendium index was also suitable for establishing a comprehensive computer-loaded data base on the subject. 2. To develop a comprehensive system of guidelines to cover salient aspects of analysis, selection and design of electrical, mechanical and structural components of a hoisting system with a view to establish parameters for maximum safety and efficiency of operation, installation and revision of the system. 3. Considerations in the mine shaft layout, design, sinking, outfitting, testing and commissioning are set forth against the background of requirements of applicable parts of Code of Federal Regulations, Title 30.

Jan 1, 1988

By Sam Palombo, James José Varela

"Washing can be defined as the process used to eliminate or reduce the dirt (presence of fines or slimes) of one material. It is also the basic and oldest process applied to ore preparation for further concentration. This process, in combination with commnution and screening, form the so called washing plant and in many cases produce end products. This work describes the factors that influence the performance of the washing process and main washing technologies used in the mineral processing industry.The Hydro-Clean represents, since 1998, a state of the art technology in this area and is comprised of one of the most innovative solutions developed at HAVER & BOECKER. HAVER & BOECKER design and manufacture technology for mining and related industries. Today the Hydro-Clean is manufactured and sold throughout the Haver Screening Group (HSG) which is comprised of HAVER & BOECKER Maschinenfabrik Münster, HAVER & BOECKER Latinoamericana – Brazil and W.S. Tyler Canada.The Hydro-Clean is a washing technology based on utilization of high pressure water jets. To follow we will discuss how the Hydro-Clean’s different modes of operation result in several advantages related with mine waste reduction, recovery of saleable product, lower capital costs, superior washing results, less stress for the material, a wider range of application and less maintenance cost. Aside from these positive factors, its lower operation costs related to the consumption of energy and water are presented. It is important to note that the quantities, tonnages, data and management concepts presented in the paper are accurate for the time of writing and are subject to change and modification as operations progress."

Jan 1, 2009

By Rudolf E. Greuer

Two computer programs for the real-time calculation of concentration distributions in the ventilation system of underground mines were designed. They allow one to simulate the movement of fumes or other contaminants with the ventilating air, the calculation of concentrations at any given point and time, and the determination of the total contaminant exposure at different places in the mine. Multiple and time variable contaminant sources can be handled and recirculation is no obstacle. The, first program is based on an airflow distribution which has, in the input data, to be stated. The second program has been linked to in existing program for the calculation of airflow distributions. Two versions of this second program are provided to accommodate the two existing programs. Several executed examples demonstrate the program performance.

Jan 1, 1981