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By Terry J. Briggs, Ian J. Kelso

"Abstract- The Century mine in northwest Queensland,Australia, is a major sediment-hosted zinclead- silver orebody, located within the western Mount Isa Basin. Stratiform, carbonaceous shalehosted mineralization is developed within Proterozoic shale and siltstone sequences of the Lawn Hill Formation. CRA Exploration Pty Ltd (CRAE) discovered the Century deposit in 1990 and open pit mine development commenced in 1998 under the ownership of Pasminco Century Mine Limited (PCML). The deposit contains an identified mineral resource of 105 Mt @ 12.1% Zn + 1.7% Pb + 46 g/t Ag (Broadbent and Waltho, 1998) which is currently being mined at a rate of 5 Mt/y.In December 1998, an exothermic reaction was observed between explosives spilt onto shale drill cuttings from a blasthole. This reaction is referred to as ‘reactive ground’ and has the potential to cause premature detonation of explosives in a charged blasthole with catastrophic consequences. The incident led to a subsequent geological and geochemical assessment of the different rock types at the Century deposit and their potential reactivity with ammonium nitrate-based (AN) explosive products. This assessment was conducted in an attempt to characterize zones safe for blasting with uninhibited AN-based explosive products. A direct result of this study is a set of procedures outlining the use of different AN-based explosives in areas defined as potentially reactive or non-reactive for the different lithologies present. Black shales contained in both the waste and mineralized sequences contain the sulfide mineral pyrite (FeS2), which is thermodynamically unstable with AN contained in commonly used explosives. The potential reactivity of partially weathered pyrite can ultimately result in the decomposition of an AN-based explosive. It was determined that only pyritic black shales showed any potential to react with AN, irrespective of their position in the stratigraphic sequence. Other lithologies present at Century, including Proterozoic sandstone and Cambrian limestone, are classified as non-reactive with AN. This is a geological risk management situation present at the Century mine."

Jan 1, 2001

By J. Petersen, T. Moyo, D. Maharaj

This study investigates the effect of ammonium thiosulfate on the leaching of gold from the contacts of unpopulated cut electronic printed circuit boards (PCBs). Studies in the literature show ammonium thiosulfate to be a viable option in comparison to the commonly used but hazardous cyanide leaching route. However, the simultaneous dissolution of Cu interferes with the Au leaching reaction, often limiting the extraction to moderate levels. To optimize the system, an ammoniumthiosulfate system containing 0.5M ammonia, 0.5M ammonium thiosulfate, and Cu(II) sulfate at varying concentrations was studied as an alternative hydrometallurgical option for the recovery of gold from PCB. Tests were done with unpopulated standardized model PCBs (cut to 15 x 30mm pieces). Gold extraction of 47.3% was achieved after a 24-hour leaching experiment with a 0.02M copper background concentration. Further experiments focused using different initial background Cu concentrations (0M – 0.1M) to determine its effect on both gold and copper extractions. An initial Cu concentration of 0.045M resulted in the highest gold extraction, with 93.7% extraction after 6 hours of leaching. Further analysis of the data shows that the kinetic rate of gold leaching is correlated in a first-order dependence with the actual copper concentration in the solution, which increases with the concomitant dissolution of copper and gold from the PCBs. A lower threshold of 0.017M and an upper threshold of 0.15M copper in solution were determined, in between which gold extraction shows this first-order rate increase. The upper threshold is potentially linked to the onset of copper precipitation, which blinds the metal surfaces, slowing further leaching. The findings of this study suggest that in an industrial recovery operation of gold from waste PCB using the thiosulfate method, copper concentrations need to be carefully monitored, and the process should preferably operate with uncut or coarsely cut untreated boards to minimize exposure to copper circuits.

Dec 11, 2024

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

Jan 1, 1994

By William G. Pariseau

"This contribution describes linkage of stress concentration at the face of tabular excavations, for example, longwall coal mine panels, at a meter scale to surface subsidence at a kilometer scale. Joints are considered according to orientation and spacing at an intermediate scale. A new dual-node, dual-mesh technique based on first principles uses a multi-scale technique that allows for computing whole mine surface subsidence while also allowing for computational details of stress, strain and displacement about main entries and pillars, panel entries, and bleeder entries. The technique is implemented in a conventional continuous Galerkin finite element program that has been in use for many years.The advantage of the dual-node, dual-mesh technique is the capability to do a whole mine analysis that takes into account interaction between all sections of a mine where stress transfer over long distances is often the rule rather than the exception while at the same time allowing for details of stress concentration at working faces where caving is initiated and subsidence begins. The technique, outlined recently during ARMA 2014 (Pariseau, 2014), is applied to an underground coal mine in central Utah where mines are developed from outcrops in steep canyon walls. Meshes contain over 10 million, three-dimensional nodes; runtimes are typically overnight, as a practical matter. Results compare favorably with mine subsidence measurements made over a period of several years and show the capability of this new technique to provide useful design guidance for tabular deposits such as underground coal mines. The role of variability or uncertainty in strata properties is illustrated in element safety factor distributions about development, bleeder and main entries and pillars. Such details are new in this contribution. Energy tracking for mine-induced seismicity is also new. Results show a high correlation between number of events and failed elements and suggest a simple index to mine induced seismicity."

Jan 1, 2015

By O N. Savassi

A rigorous methodology for prediction of flotation plant performance is presented, which is based on a standardised batch flotation test called the MFT. The work is part of the FLEET program, developed by MinnovEX with the support of seven large mining firms. The methodology is divided into three key steps: direct measurement of the kinetics of true flotation in the pulp, independent of hydraulic entrainment and other froth effects in the batch cell; benchmarking of the industrial plant over a broad range of ore types and operation conditions; and model calibration to describe froth effects in the plant, with a minimal number of parameters. To date, the methodology has been applied with success at several industrial plants, encompassing copper, gold, nickel, iron, lead, zinc and coal operations. This paper includes a description of the laboratory testwork procedure and scale-up to full industrial circuits. Examples of industrial application are also presented.

Jan 1, 2005

By S. Ramachandran, R. G. Woolery

INTRODUCTION Union Carbide is currently operating an in-situ leach project on the Palangana Dome area in Duval county. This deposit meets all the requirements for in-situ leach in that the ore (1) is below the water table, (2) is in a permeable horizon, (3) is amenable to chemical leaching, and (4) is confined by impervious layers. This project has been under commercial production since 1976, and its capacity has been expanded on three occasions since going on-stream. Recently, additional uranium reserves were discovered on the Rogers-Cardenas (R-C) property about 32 km north of the Palangana operation. The ore is located within the Oakville sands and its characteristics are quite similar to those of the ore at Palangana. Both are an unconsolidated Arkosic sand high in clay and calcium carbonate. The R-C ore, however, is somewhat coarser with a mean particle size of 0.15 mm as compared to a mean particle size of 0.07 mm for the Palangana ore. In all respects it would appear that this ore would be a candidate for in-situ leach as a satellite operation to Palangana. Unfortunately, R-C ore is above the water table and, therefore, not amenable to the Palangana practice. Because of the limited known reserves in this deposit, it is readily apparent that conventional mining and milling are out of the question. However, because of its proximity to our Palangana operation, it seemed worthwhile to consider other options. The most viable route based on our past experience was to heap leach the ore. Our recent success at our Gas Hills facility and our Maybell operation, in employing a heap leach practice to our marginal reserves seemed to be a logical approach for processing this ore. Our experiences at both locations are described in "Heap Leaching - A Case History" by R. G. Woolery et al., Mining Engineering, March 1978. In both instances the process is an acid leach circuit and acid consumption averages 20 kg/t H2SO4. A preliminary feasibility study showed that because of the high strip ratio required for the R-C project to be successful, additional ore reserves must be located and that a method of heap leaching with an alkaline circuit would have to be developed. As a result of this paper study, the decision was made to proceed with a program of additional exploration drilling to determine the total ore reserves that could be mined economically. The Mining Department will evaluate each ore zone for cutoff grade, strip ratio, and expected mining cost. At the same time, a laboratory program to evaluate the available core samples for amenability to heap leaching with respect to an estimate of uranium recovery and processing costs was developed. This program is currently in progress, and at this time, we are just completing our process amenability study. BENCH-SCALE EVALUATION OF THE R-C ORE The initial bench-scale slurry leach tests on the R-C ore showed an acid consumption in excess of 200 kg/t H2SO4. These data, of course, discouraged us from considering this process route. Not only would the acid cost be prohibitive, but the gypsum generated by the reaction of the sulfuric acid with the calcium carbonate of the ore would severely effect the percolation of the lixivant. For this reason, the laboratory program was directed toward an alkaline circuit compatible with heap leaching. Because of the proximity of the R-C property to our Palangana operation, it seemed advisable to integrate the processing of this ore into the production at Palangana. Doing so would enable us to bring the R-C property into production by merely enlarging our present facilities at Palangana; otherwise, construction of a grass roots plant would be necessary. Ideally, the simplest method would be to construct the heaps at Palangana and employ an ammonium carbonate/bicarbonate leachant compatible with the in-situ production liquor. The product liquors could then be co-mingled or processed separately as desired. To determine if this goal was practical, samples of the R-C ore were obtained, and a laboratory program initiated. Heap leach amenability testing consisted of preliminary bench-scale evaluation to determine optimum solution strength and ultimate uranium recovery, followed by small column tests to confirm the bench-scale metallurgy and to determine percolation characteristics. These bench-scale tests are being followed by pilot-scale testing approximating field conditions. As expected, the bench-scale tests showed that the dissolution rate is considerably slower for alkaline leach than has been our experience in acid leaching. Because of the slower reaction rates, product liquor grades will be lower than for acid, as greater volumes of solution are required for satisfactory uranium extractions. The greatest influence on reaction times found in the laboratory was the carbonate/bicarbonate strength and oxidant addition. However, the higher salt concentration reduced the efficiency of the IX resin circuit and about 25g/L salt proved to an upper limit compatible with subsequent IX treatment. The oxidant contributed significantly to the early extraction rate but seemed to have only minimal effect on the total practical U308 extracted or the time required to achieve it. This variable will require larger scale testing to determine if the added cost of the oxidant is actually justifiable. Thus, the small-scale laboratory slurry tests, based on the 0.088% U308 sample available, indicate that leaching at 25g/L ammonium carbonate/bicarbonate, with or without oxidant, we might expect an 80-85% U308 extraction on this ore.

Jan 1, 1979

By Abid O. Adekunte, David Greentree, David Hilton

"The construction industry is witnessing an upsurge in demand by property developers to have basements incorporated into their residential and commercial development schemes. This is attributable to perceived improvement to aesthetic and commercial values of buildings by underground space, amongst property owners. This trend is compelling more owners of buildings with no basements to explore the possibility of creating underground additional space beneath their existing buildings. The creation of a basement beneath an existing building entails the installation of a reinforced concrete (RC) underpinning system, temporary works, bulk excavation to basement formation level and construction of the basement structure, with associated aesthetic finishes.The L-shaped segmental RC underpins are a common system of choice for such underpinning works. They are multi-functional; providing vertical support to the existing superstructure, providing lateral support to earth and groundwater, supporting temporary and permanent surcharge loads from adjacent structures and services, whilst also forming part of the basement slab. This paper focuses on the first function of the Lshaped RC underpins; capacity under vertical compressive loading from the superstructure. The most common industry approach for the prediction of vertical capacities of L-shaped underpins is the oldfashioned Meyerhof’s (1953, 1963) effective width method, which is based on the traditional bearing capacity theory, whilst accounting for correction factors for eccentricity and foundation depth.The traditional bearing capacity formula is surrounded by a significant degree of uncertainty; estimates are highly user-dependent, as capacity prediction is sensitive to the chosen angle of shearing resistance (?), while there are difficulties in accurately measuring the in-situ ??value of cohesionless soils. In addition, the effective width approach has been shown to be very conservative by many research workers. The segmental RC underpins presented in this paper have been designed with an alternative less conservative method, which involves the use of a recently proposed consistent work-hardening plasticity model calibrated with field-based tests on footings of various aspect ratios and relative embedment in sand. The paper highlights the principles behind underpinning works for urban basement construction, the design of L-shaped RC underpins with the recently proposed plasticity model and the advantages of the method over traditional bearing capacity solutions and pre-existing plasticity models."

Jan 1, 2016

By Joe E. Bryan, John P. McDonnell

"Underground mine rib stability is an ongoing safety issue in all mining applications. Rib-related accidents present at least as many problems as roof-related ground control. Underground coal mine rib safety is continually emphasized as a primary safety concern at every mining operation. The application of rib support materials is increasing as more difficult mining conditions are experienced due to deeper cover, more variable depositional environments, thicker seam heights, and multi-seam mining effects. As the coal seams become thicker, the need for rib control increases. A variety of rib control methodologies has been documented and research continues on mine design methods and support techniques to reduce or eliminate rib-related accidents.This paper presents an alternative rib bolting technology using proven conventional rib support systems that allows mine personnel to more efficiently install rib support during development operations. This alternative rib bolting method allows for additional rib support with a single support installation.INTRODUCTIONRib control continues to garner the watchful eye of both concerned operators and government officials due to many factors: inherent safety within entries and crosscuts, actual rib stability history, safety improvement, cost control, improved mining methods, and maintaining production goals. As demonstrated through many years of actual practice, desired outcomes can be achieved by scrutiny, innovation, and combining improved practice methodologies.Rib surface control continues to be a necessary part of the production process. However, production and safety issues persist at the point of installation using current control apparatuses and methods to install rib controls. Among others, current issues of rib control are quantity of bolts required, overall time required for installation, ease of apparatus installation, flexibility of the system for appendage installation and operator exposure. A need exists to decrease the time miners spend installing rib control apparatuses, while increasing the safety of such installation processes. At present, installing rib control devices (primary rib bolts, plates, mesh, and, in most cases, secondary rib bolts and plates) is performed in a very congested area and is time-consuming process at the primary coal-winning area: the face. The MatLok system addresses these problems."

Jan 1, 2015

The current genesis of pisolites and channel iron deposits (CID) has been defined in a way that presents CID as unique in the world. The published material on detrital pisolites mostly details the West Australian CID; most other global occurrences of iron-rich pisolites would seem to have an in situ genesis. An alternate theory for the origin of the Hamersley CID is suggested, involving an older greensand protore derived from volcanic ash and deposited on a glaciated Hamersley surface. This paper presents the author's observations, sampling and mapping after several years study and exploration for pisolites and CID in WA and in some locations overseas. It is not what one would call scientifically complete, and possibly asks more questions than it answers, but it is presented as a point for discussion.

Jan 1, 2009

By D J. Williams

Tailings management is typically constrained by many factors that are taken as given. These include the tailings production rate that must be accommodated, the desire to deliver the tailings cheaply and easily as a slurry using centrifugal pumps, and a tailings footprint limited by the desire to minimise capital costs leading to a rapid tailings rate of rise and a low settled density. Other constraints include wall raises being delayed due to limited capital, inadequate tailings water and rainfall run-off management, and the net present value (NPV) approach to capital expenditure that puts off expenditure at the expense of performance. All of these constraints result in the tailings occupying much more volume than is necessary, being difficult to rehabilitate, and not being capable of supporting a beneficial post-mining land use or function. Ultimately, such constrained tailings storage facilities will assume very large dimensions, but each will typically be developed as a series of smaller storages that are filled rapidly and added to progressively. The optimal stage to efficiently recover process water is in the process plant, but this is limited by the desire to use centrifugal pumps rather than the more costly positive displacement pumps, which also require greater management of tailings deposition. Process water can also be recovered from the tailings storage facility (TSF) decant. Tailings water that is not recovered from the plant or the decant will be lost to entrainment within the tailings, seepage to the foundation or through the wall, or evaporation. However, tailings water is not always suitable for reuse in the plant, particularly when it becomes contaminated by the processing, and it may simply be stored on the TSF. This will exacerbate the seepage of contaminated water to the environment, and will limit the densification of the tailings that would otherwise occur due to desiccation of the tailings on exposure. If a larger tailings footprint were developed early, cycling of tailings deposition in thin lifts between a number of cells could occur, maintaining the dormant cells dry to enhance desiccation and densification, and ultimate rehabilitation. The end result would be a much reduced stored volume of tailings, the potential for upstream raising, and ease of rehabilitation to a high level. The higher initial capital cost would be overtaken by reduced ongoing capital, operating and closure costs.CITATION:Williams, D J, ?2014. An alternative whole-of-life approach to tailings management, in Proceedings Life-of-Mine 2014 , pp 285–298 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By C. Clark, R. Pakalnis, B. Seymour, L. Martin, M. Stepan, M. Roworth, C. Caceres

The National Institute for Occupational Safety and Health (NIOSH), Office of Mine Safety and Health, Ground Control Engineering Branch is investigating the use of shotcrete in weak rock mass mines with the objective of reducing fatalities and injuries resulting from rock fall accidents. When shotcrete is used as part of a multielement ground support system, there is a need to know the support characteristics that the shotcrete contributes to the overall system. To quantify the support provided by the shotcrete, flexural strength tests were conducted with two common, commercially available fiber-reinforced shotcrete (FRS) mixes using the round determinant panel (RDP) test method, ASTM C 1550. A portable RDP test machine developed by NIOSH researchers was used to determine the peak flexural load and residual load capacity for a steel FRS and a synthetic poly FRS that were sprayed using dry mix procedures and equipment. Besides the flexural strength and loading behavior determined from these tests, a method was also developed for measuring the width of cracks exposed on the underside of a shotcrete panel during a RDP test and relating these measurements to residual load values. Quantifying the peak flexural load and residual loads of a shotcrete mix through on site testing at a mine and visually assessing the loading cycle and load-carrying capability of the shotcrete applied to underground workings will improve mine safety by providing a better assessment of the stability of shotcrete supported entries.

Jan 1, 2010

By Ronaldo Luna, Donald Deardorff

"The use of load and resistance factor design (LRFD) is well documented since the 1960s for structural applications when it was first adopted into the American Concrete Institute Building Code in 1963 and was established for geotechnical applications by the Danish Geotechnical Institute about the same time under the name of Limit State Design (LSD).The Federal Highway Administration required all federally funded projects to use the LRFD method by 2007. Due to this funding requirement, recent LRFD research for geotechnical applications has focused on common bridge foundations (i.e., driven or drilled shafts), resulting in good documentation for design using these methods. Unfortunately, the practicing geotechnical engineer does not intuitively know how to create resistance factors for deep foundations systems that have design methodologies and installation techniques significantly different from driven or drilled shaft foundation systems.Helical anchors typically use semiempirical methods for preliminary capacity design and a torque correlation methodology during installation to verify required capacity. Here we provide basic information on the LRFD methodology for helical anchor design. Additionally, the First- Order Second Moment (FOSM) method is applied to a previous helical anchor study (Hoyt and Clemence 1989) to determine resistance factors for an ultimate tension capacity strength limit state.MethodologyThe concept of the LRFD method is to establish factors for both the load and resistance that account for uncertainties using probability theory. The factors increase the nominal load and/or decrease the nominal resistance such that the probability of failure is an acceptable value. LRFD also accounts for the effects of load combinations (dead, live, seismic, wind, etc.) by determining the probability of the combined occurrence. The general form of the LRFD equation used by American Association of State Highway and Transportation Officials (AASHTO) is:"

Jan 1, 2009

By L. M. English

In the past 30 years, the research arena has changed significantly in all fields, mining included. The relative representation of the topics of research, how the research is performed, and who has performed the research have all changed significantly over these decades. Although there are a lot of subjective ideas concerning these changes, and there is a fair amount of anecdotal "analysis," there has been no actual study of these trends to verify the facts. The period studied is divided into six five-year periods for the purposes of comparison. Some of the trends analyzed are as follows: ? Co-authorship ? Research areas ? Research sponsors ? Cooperative efforts

Jan 1, 1995

By S. M. Moore

Each year more than 400 coal miners are injured (fatal and non-fatal) by rock falling from between or around roof supports. Many of these injuries could be prevented by the installation of roof screen (wire mesh). However, many coal mines may be reluctant to use roof screen because of the added cost. The goal of this study was to determine the potential saving in workers? compensation (WC) premiums that could be achieved due to a reduction in rock fall injuries after roof screening. The WC rate-setting methods utilized by IL and KY were investigated in this study. Using actual data obtained from MSHA, national and state WC bodies and individual insurance companies (e.g. average cost per injury, loss cost rate, number of injuries per year, number of injuries preventable each year with roof screening), hypothetical mines (representing two mine sizes: 67 and 150 employees) were constructed with realistic ranges for estimates of injuries and WC premium costs. Using each state?s actual WC rate-setting formulas, total savings in WC costs after a three-year period were determined. Across both states, savings in WC premiums ranged from 1.8% to 14.6% when injuries were reduced by 10% to 20%. Additionally, an economic analysis was performed for an existing mine consisting of 67 employees. For this mine, the annual cost of a roof screening program was estimated at $240,000. If the screening prevented 13 ?struck by? WC claims over a 3-year period at this mine, the reduction in WC premiums alone could pay for the entire screening program.

Jan 1, 2010

By K. Barron

"A computer program is described which will analyze the stability of coal pillars. The extent of the peripheral failed zone and the central elastic core of the pillar, the stress distribution across the pillar and its strength can be determined by a ""sliceby- slice"" analysis technique. The program incorporates a brittle/pseudo-ductile transition point so that a pillar may fail by brittle fracture or pseudo-ductile yield. It is shown that:(i) There is a minimum pillar stress below which fracture cannot initiate; hence there is a minimum pillar strength. (ii) In specific circumstances (narrow pillars) , once fracture initiates at the pillar edge it must propagate immediately to the pillar centre. This unstable situation is called ""catastrophic"" failure and might be regarded as a model of a pillar ""burst "". (iii) For wider pillars, the failed zone ex tends gradually into the pillar as the loading increases until it reaches the pillar centre when the pillar is deemed to have f ailed. This is called the point of ""ultimate"" failure.The critical pillar widths to avoid catastrophic and ultimate failure can be calculated.A hypothetical analysis of five rib pillars is given to illustrate the application and the results obtained from the model. The model is then used to analyze three case histories taken from the literature. The first demonstrates the applicability of the brittle/pseudo-ductile failure criteria for Pittsburgh coal. The second compares the pillar strength formulae derived here with those of Salamon and Munro(13) and Bieniawskiou . The final case history compares the model analysis with field stress measurements in a yielding pillar. In all cases the model results are compatible with the case histories."

Jan 1, 1984

By Noel Brien, O&apos

General: National Asbestos Mines Limited has operated an asbestos mine and mill near Thetford Mines, Quebec, since 1958. As a result of additional exploration of the orebody in 1962-63, it became apparent that long-range mining plans would have to be revised to allow for a large increase in both ore and waste tonnages. It was decided that the task of checking out possible mine designs could best be analyzed by a computer to obtain data from which an optimum open-pit configuration could be plotted. Object: It was desired to arrive at an optimum open-pit configuration with an over-all pit slope of 45 degrees and a minimum operational width of 200 ft. at the bottom. Calculations were based on assay values from diamond drill holes in cross sections of the mine. The computer's assignment was to select and describe, by coordinates, the optimum cross section for the mine. Mine engineers then assembled the various optimum sections to produce the pit outline. As this would not necessarily result in a practical mining arrangement, the computer also provided alternate solutions,

Jan 1, 1965

By C. Allen, H. Zhang, L. Falk

Shock loss occurs when there is a change in the area of a drift or airflow direction. These shock losses accumulate throughout the ventilation system and will increase the mine resistance as underground development deepens or becomes more complex. Hence, shock losses should be carefully considered when establishing the resistance through a series of branches as they can be a significant contributor to the estimated mine resistance. Using computer software to simulate a ventilation system is common but applying reasonable shock loss factors in the software is not a simple or direct process. Determining shock loss values includes the consideration of historical measurements, application of factors and formulas from literature based on the geometry and complexity of junctions, and the geometry variation at different locations. For example, the shock loss at the bottom of a ventilation raise can consist of a 90° bend and a sudden area change. This paper presents an approach for estimating shock loss factors near raise junctions using computational fluid dynamics (CFD). Current and future work will also be compared with field measurements. Results from this approach are compared with software defaults as well as classical techniques in published literature and standards.

Sep 1, 2024

By B. Tolwinski

Long-Term Open Pit Scheduler or L-TOPS is a new software package for evaluating, planning, and design of open pit mines. L-TOPS transforms geological, economic, and technical information about a mineral deposit into a detailed mining schedule that maximizes economic efficiency without neglecting the practical requirements of mining operations. L-TOPS has been customized for integration with Placer Dome Inc.'s existing in-house resource/reserve optimization software to function as a primary tool for property acquisition studies and for the designing of mines for feasibility studies. This pa-per presents an overview of L-TOPS and reports on its application to Placer Pacific's (Placer Dome's Australian affiliate) Mt. Rawdon project.

Jan 1, 1996

By Marcos Paulo da Silveira, Carolina Conter Elgert, Messias Prado de Almeida, Rafael Stella Galdino

Microalloyed steels have been intensely studied as an alternative to increase strength by grain refinement and/or precipitation hardening. Long steel products have a wide range of opportunities to apply well-defined principles of controlled rolling (CR) in order to increase strength and toughness with reduced costs. To evaluate the effect of microalloying elements as Nb and Ti with thermomechanical treatment, three conditions were tested: DIN 16MnCrS5 with hot rolling (HR), DIN 16MnCrS5 with Nb and CR and DIN 16MnCrS5 with Nb+Ti and CR. Austenitic grain size, obtained by direct quenching after rolling, has shown a decrease 32% with Nb addition, and 29% when Nb and Ti was added. Controlled rolling generated higher grain refinement and pancacked grains near the surface. When comparing ferritic grain size, controlled rolling plus niobium and titanium additions resulted in significant reduction of 35%. Yield strength increased 15% with microalloying and CR and more than 270% in impact toughness, even at -40°C.

Oct 1, 2019

By D J. Thobaven

Project financing challenges caused by declining commodity prices have heightened the need to improve project metrics such as minimising capital payback and maximising net present value (NPV). Khnaiguiyah Mining Company (KMC) recognised the need to improve early cash flows at their Khnaiguiyah zinc-copper project after initial studies showed the project could not provide the desired financial outcomes necessary to compete for investment dollars in the current market.A significant improvement to the project cash flow in early years was realised through simultaneous optimisation of cut-off grade, mining sequence and strategic stockpiling. This improvement was achieved by accelerating mining in the pits, stockpiling lower value material and feeding higher value material directly to the plant. The schedule required higher mining and stockpile capacities than the sequential mine plan although minimal change occurred in the definition of ore and waste as measured by tonnes and grade. The practicality of the strategic schedule was demonstrated by implementation in the production plan with excellent conformance.CITATION:Thobaven, D J, 2014. An application of simultaneous optimisation – improving the payback period of a zinc-copper open cut project, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014 , pp 137–144 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 24, 2014