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By D Kentwell

For deposits where a full assay suite of elements/oxides is available and where several variables are critical to the evaluation of the deposit, it is desirable for the estimated block values of all elements/oxides to sum to 100 per cent of the weight per cent. This is difficult to achieve in practice and the estimation methods used need to be designed to minimise bias of the key elements/oxides by ensuring that block estimates are consistent with the individual assays and that the estimated totals are as close as possible to 100 per cent.This paper presents a case study of two domains within a magnetite hematite deposit, one with a high level of drilling classified as Measured and one with a lower density of drilling classified as Indicated. Several estimation methods were tested to evaluate their ability to maintain consistency of the proportions of each element/oxide and sum to 100 per cent. The following methods were evaluated: ordinary kriging, co-kriging, inverse distance and min/max autocorrelation factors (MAF) all as implemented in the Geovariances Isatis software.The comparisons showed that the MAF method produced the most robust estimates in terms of block totals and that grades and tonnages at the reporting cut-off were not significantly different between methods. At an individual block scale all methods produce artefacts that are less than desirable but can only be remedied by additional drilling.CITATION:Kentwell, D, 2015. Comparison of min/max autocorrelation factors, kriging, co-kriging and inverse distance block totals – a magnetite hematite case study, in Proceedings Iron Ore 2015, pp 465–474 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By C Taylor, E Retz

Sequential Gaussian conditional simulation of wide-spaced exploration data has been completed in order to optimise a drilling grid at Fortescue’s Eliwana deposit. The aim of this drill grid is to improve the confidence in the Inferred Mineral Resource. A mineralised domain within the Joffre Member of the Brockman Iron Formation was selected to be tested as it makes up a large proportion of the Inferred Mineral Resource and is drilled at a reasonable spacing.One hundred realisations of Fe, SiO2 and Al2O3 grades were completed in the domain. Three realisations were selected for each analyte to represent the range of equiprobable values and a further realisation was selected to represent the truth data set. These realisations were used to flag theoretical drill holes on varying spacings and patterns. The theoretical drill holes were used to estimate into block models using ordinary kriging.These estimates were analysed and the accuracy and precision of each drill spacing was assessed. It was concluded that the greatest improvement in estimation results was noted at either a 100 m × 50 m or 50 m × 50 m spacing.CITATION:Taylor, C and Retz, E, 2015. How much do we need to drill?, in Proceedings Iron Ore 2015, pp 491–500 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By A Fowler

The unpredictability of the location, grade and width of narrow-vein deposits presents difficulties for mine geologists and engineers attempting to estimate, schedule, forecast and mine their deposit profitably. This unpredictability stems not only from the grade distribution within a vein, but also the irregularity of vein geometry in three-dimensional (3D) space. Traditional deterministic methods of modelling vein geometry do not account for the inherent variability of the veins, resulting in significant errors when using limited drill hole information. This paper presents a probabilistic method of modelling narrow-vein deposits using sequential gaussian simulation (SGS), which can quantify this variability and improve production forecasts. Vein geometry, thickness and grade are simulated to provide a number of different equi-probable realisations that each honour the drill hole and face sample input statistics. An example is presented from the Augusta narrow-vein Au-Sb mine in central Victoria, where the simulation results are validated against the input data and survey pick-ups of the veins along ore drives.

Aug 22, 2011

By C Muñoz, N Ortega, J Vergara, D Avilés

Mass blasts have been the subject of extensive studies because of their potential benefits in mining operations. However, it has not been possible to advance due to the paradigm that states that the blasts near the designed final wall or to critical structures have to be low tonnage. Another reason is the absence of specialised simulation programs to more precisely determine the influence of the energy released by blasting to the rock mass. There are some studies that demonstrate the possibility of massive blasting that maintain the design of the final slope (Farias and Santander, 2014). It should be noted that these studies develop the concept that the size of the blast should not increase the peak vibration in the rock mass.The study was carried out in partnership with Los Colorados Mining Company. Analysis, reconciliation and wall control for two types of blasts were performed:blasting production: approximately 850 000 t, with special wall control considerationsblasting of approximately 950 000 t over the entire width of the bench.Monitoring systems were used in the near field and far field using examination of holes in order to obtain information to assess the impact on the rock mass and then to compare these to conventional blasting combining production and contour blasting.The results show that for a large blasted volume, production and/or those generated from toe to crest on the bench did not produce more impact in the rock mass than the conventional blasting.The conclusions drawn from this study are:an improvement in the excavation rate of 7500 t/h that is higher than the maximum rates normally obtainedthe peak vibrations of the P wave generated by both blasts did not exceed the peak particle velocity criteria (Vergara and Muñoz, 2014)holes examination indicating no difference in their entire length, in pre- and post-blast reviewover US$2 000 000 savings generated in the year (estimated by the client) (Avilés, 2014).CITATION:Vergara, J, Muñoz, C, Ortega, N and Avilés, D, 2015. Benefits related to applications of a mass blast in open cut mining, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 625–632 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By S Coward, C Xu, P A. Dowd

"This paper provides an assessment of the current challenges in strategic mine planning and design and suggested approaches for addressing them. The specific challenges covered are: realistic quantification of downstream processes applied to orebody models to provide an integrated approach to mine design and optimisation modelling, estimation and simulation of geometallurgical variables and their integration into resource and reserve estimation and mine planning modelling, estimation and simulation of new variables for new forms of mining – deep mining, particularly block caving, and solution mining flexibility in planning and design to manage risk and minimise its impact information technology (IT) infrastructure and platforms for rapid online data collection, storage, access and processing.Most of these challenges require new types of data, variables, modelling and estimation methods. Foremost among the new types of variables and data are geometallurgical and dynamic rock mass characterisation variables. New types of data and data collection include rapid generation of very large amounts of online sensor data and the consequent need for rapid processing and modelling of these data.This paper outlines the challenges and strategies in each of these areas and uses examples of models and outputs to illustrate approaches and potential solutions.CITATION:Dowd, P A, Xu, C and Coward, S, 2014. Strategies for mine planning and design, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014, pp 41–50 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Nov 24, 2014

By M Salu

Ventilation control devices (VCD’s) and bulkheads are critical controls used for managing risk in underground coalmines. Following the Moura No. 2 mine explosion in 1994, the Queensland Mines Department required that the design of VCDs be based on live testing (such as in a mine testing explosion gallery) and more recently on numerical modelling using measured seal material properties. While the stipulated methods represent an idealistic solution, they are not practical to implement as solutions to daily challenges presented in underground coalmines.Not all VCD’s and bulkheads are placed in areas which have been mined to standard roadway dimensions. Operationally, VCD’s may be required to be placed close to intersections or require large dimensionally different areas (eg overcasts and pre-driven recovery road shutters) compared to the live testing or numerical modelling originally undertaken.This paper describes the various structural engineering methods utilised to account for the design of VCD’s where the operational requirements are non-standard. Siting and construction of VCD’s near the entry of cut throughs and geological structures requires additional design considerations. While VCDs constructed from a single material (eg sprayed mining plaster) are relatively straightforward to design, the addition of multiple openings (doors) or large pipes will complicate the structuralanalysis significantly.Temporary bulkheads used for the containment of grout such as for the construction of pre-driven recovery roads in coal mining have a significantly different risk profile to VCDs and should be designed accordingly. Recent failures and near-misses in such bulkheads have highlighted this issuefor the authors.The existence of an Australian Standard for Underground Mine VCDs and Bulkheads would go a long way to addressing many of the issues raised in this paper. Without a single, industry-accepted reference for VCD design, individual mines will have to continueto exercise their best judgements using whatever local expertise is available to them.

Oct 12, 2022

By B Reynolds

The process design for green- or brown-fields metallurgical plants typically starts with two fundamental elements: the detailed mass and energy balance and the plant flow availability. The detailed mass and energy balance is generally developed using process flow-sheet simulation software and represents the steady-state operation of the plant at normal operating conditions. The plant flow availability, also known as a range of other industry-specific terms such as ‘flow availability’ and ‘operating factor’, describes the ratio of plant production to the steady-state mass balance. These elements are then used together as the basis for equipment selection and design.Plant flow availability takes into account a myriad of production loss factors including planned and breakdown maintenance, feed availability and composition changes, variability in operational performance and the interaction of batch and continuous unit operations. The combined impact of these loss factors are complex and as such plant flow availability is usually derived using a top-down approach such as comparison to an existing operation, a group of similar operations or to an anecdotal ‘industry leading’ benchmark. These top-down approaches are fraught with risk however, as they cannot factor in changes to equipment selection, flow-sheet or plant operating philosophy, and can result in over-expenditure of scarce capital or shortfalls in plant production.The use of dynamic simulation to provide a bottom-up alternative to quantifying the plant flow availability in metallurgical plant design has been growing rapidly in recent years. It provides a technique to quantify the combined impact of the loss factors described here and a tool to support process design and equipment selection by facilitating a range of sensitivity and what-if analyses. This paper presents a case study of how dynamic simulation has recently been used to support the

Aug 8, 2011

By Hoskings R

Conceptual studies of the run of mine stockpiling and reclaiming requirements com- menced in 1975. Final selection of equipment was made in 1978 and the equipment was commissioned in June, 1979. Equipment selection was made after extensive investigation and visits to existing facilities throughout the world. The equipment selected is described and com- pared with the alternatives. Foundations were designed to minimise clean up effort and to ensure stability of equipment under design wind loading without storm bays. Program- mable logic control enables the stacking and reclaiming facility to provide a blended feed to the coal preparation plant.

Jan 1, 1980

By D Reid

Characterisation of non-plastic segregating tailings presents a number of challenges: the material cannot typically be sampled in an undisturbed state, remoulded samples exhibit different behaviour to undisturbed samples, penetration test interpretation methods suffer from uncertainty in their application to many tailings and the tailings gradation changes considerably over relatively small lateral distances and depth. In addition, the process liquor of many Western Australian tailings contains large quantities of salt and other dissolved solids, which can further complicate laboratory test programs. To provide an insight into the variability at a typical gold tailings facility, four gradations of the material were tested to obtain critical state parameters. Historic surficial tube samples were collated and organised by gradation to establish the range of densities typical after settling and air drying. These were then compared to laboratory-accessible densities. Finally, tube samples were tested in the laboratory to observe if the large-strain material behaviour was consistent with the inferred critical state parameters for similar gradations. Site pore fluid was used in the testing of the tube samples in an attempt to assess if the pore fluid modified the critical state conditions of the material.The tube samples tested all exhibited strongly dilative behaviour in undrained shear. Of the four samples tested, only one exhibited behaviour consistent with the inferred location of the critical state line relevant to the material’s gradation. The reasons for the unexpected behaviour of the other three samples include layering within the samples, the potential influence of large quantities of salt in the pore fluid and likely differences in the gradation of the fines components of samples taken from mid-beach and the decant area.CITATION:Reid, D, 2015. Observations on the behaviour of a gold tailings with hypersaline pore fluid, in Proceedings Tailings and Mine Waste Management for the 21st Century , pp 191–200 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 27, 2015

By S N. Donaldson, B L. Hollis, K Wynn, L Potts, M Evans

A real-time load-haul-dump (LHD) dispatch optimisation tool has been deployed at Panel Cave 1 and Panel Cave 2 at Newcrest’s Cadia Valley Operations. The tool, known as ORB, is the first known successful implementation of an optimisation-based, autonomous, real-time dispatch tool, custombuilt for underground mining operations. ORB can co-ordinate the optimal dispatch of LHD’s in a block cave to achieve a range of objectives while adhering to draw compliance and geotechnical constraints. It is an example of highly automated short-term interval control. This paper starts by outlining the pre-existing business planning process, detailing the simple logic that produced the daily draw order. Next, the motivations for a real-time dispatch tool are presented, drawing comparison to the limitations in existing planning processes and how these can be overcome through an optimisation-based, autonomous, real-time dispatch tool. Subsequently, the paper details the Ratio Bogging Draw Strategy, through which tonnes targets are expressed continuously on draw bells or draw points as a ratio of total cumulative cave draw. This signifies artificial planning boundaries are removed and that draw targets are linked within a cave; such that tonnes are unlocked by bogging other parts of the cave, instead of as a function of time. This creates incentive within the target mechanism for fixing and bogging struggling draw areas whilst allowing uncapped progress if all draw targets are progressing within compliance. Furthermore, the optimisation tool developed is outlined, including the data it uses and the objectives it optimises. Finally, detail on the improvements to operations the implementation of this tool has brought, and the business process change it has facilitated are presented using Newcrest’s Cadia Valley Operations as a case study.

Mar 15, 2021

By S Jahn, M Noel, E M. Rykaart, M Pau

Three soil cover test areas were constructed on waste rock backfilled into the Lichtenberg Pit, near Ronneburg Germany. A fourth, uncovered area was included as a control. Each test area measures 50 m by 65 m, and was constructed using full-scale equipment and construction practices. Extensive material sampling and testing was included as part of the construction. Devices were installed in each plot to monitor meteorology, infiltration, run-off, interflow, soil moisture, soil temperature, and soil pore gases. Boreholes were completed in each area and instrumented for temperature, pressure and pore gas sampling to a depth of approximately 60 m. Data from the monitoring instruments have been collected since July 2000, and have been comprehensively analysed and modelled on two occasions. The paper presents an overview of the test area design and instrumentation and the results of the two comprehensive data reviews. The reliability of the various monitoring methods is discussed and experience with several innovative methods is reviewed. The difficulties encountered in modelling the data set using common soil physical models are discussed.

Jan 1, 2003

By Y Li, S Ramazan, E Topal

A mine waste dump progression plan can provide a clear outlook of the mining operation. To produce such an output requires detailed placement schedule of the mined material, including the volume (or tonnage) and the allocated dumping location. However, current practise mainly focuses on the ore production, over-simplifying the waste material scheduling. As a result, a rock dump is often treated as a single point in long-term planning, making it difficult to predict the waste dump progression pattern over the life-of-mine. Without such a guidance, it is almost impossible to carry out progressive rehabilitation on the waste rock dumps. The lack of dumping schedule could cause delay in development construction, ie tailing storage facility and run-of-mine pad. Other downstream effect due to the over-simplification is inaccurate estimation of required truck hours, which could have huge financial impact on the operation.In this paper, mixed integer programming (MIP) models of different objective functions, ie maximise truck productivity by minimising the overall haulage distance, minimise required truck deviation between adjacent years, and a hybrid between the two objectives, are utilised to generate the long-term optimum rock placement schedules under the criteria of satisfying site specific conditions. All three MIP models are implemented in a large scale open pit mine. The numerical solutions from the models forms three different rock placement schedules, based on which, the yearly truck requirements are easily calculated and compared. The graphical results show the three corresponding waste dump progression patterns over the life-of-mine, providing the optimised long-term forecast of the operation.CITATION:Li, Y, Topal, E and Ramazan, S, 2014. Optimising the long-term mine waste dump progression and truck hour schedule in a large-scale open pit mine using mixed integer programming, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014 , pp 335–344 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 24, 2014

By E Philippa, J Vitanza, C Hawtin, B Miller

B Miller , E Philippa , C Hawtin and J Vitanza

Mar 22, 2022

By J Levett, O Rondon, I Glacken

Reporting Codes are not prescriptive on methodologies to report or classify Mineral Resource Estimation results, but the assessment of risk and uncertainty is required, and is likely to be increasingly important in future Code updates. The 2012 JORC Code indicates that relative confidence in tonnage-grade estimations should be considered for classification, and that there should be a discussion of relative accuracy and confidence levels in the Mineral Resource Estimate. These considerations are also relevant to the drill hole spacing required for various classification categories, along with other geological and data quality criteria. This paper summarises the theory, principles and current practice of various drill hole spacing analysis techniques for both achieving a desired level of resource categorisation and for optimising grade control or resource drill-out costs. The authors provide practitioners with a guide in their use to assess uncertainty and relative accuracy, as illustrated by a number of case studies. While all drill hole spacing studies rely on the assumption of consistency of the continuity model from current to future drilling, the range of methods available provide much more objective outcomes than previously adopted ‘rules-of-thumb’.

May 24, 2023

By K Fiedler, A Newell, P D. Munro

Test work is the foundation upon which the success of projects and subsequent operations is laid, reflected in Professor Arthur Taggart’s pithy observation: ‘Make your mistakes on the small scale and your profits on the larger scale’. Indeed, Professor AllanTrench has succinctly noted that ‘Get the metallurgical characteristics wrong and you don’t have a project’.With deposits becoming lower grade, more structurally and mineralogically complex, final products requiring lower levels of impurities and tailings management demanding more attention, test work now assumes even greater significance in the acceptance and successful execution of a project. This paper is the first in a series aiming to stimulate thought on the metallurgical test work process by getting back to the fundamental basics through exploring and providing some guidelines for the metallurgical testing world, noting preferred practices as well as precautions and mitigations. This covers the key characteristics of metallurgical test work programs, including what needs to be done, sample selection, quality control and associated drivers, and most importantly, selecting, assessing and managing test work facilities. CITATION:Newell, A, Munro, P D and Fiedler, K, 2018. Metallurgical test work – between a rock and a hard place, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 15–32 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By Hocking G

For some underground structures, the use of the plane strain assumption in stress analyses may not be justified. The finite element method and photoelasticity have been used to obtain complete 3-D solutions, but, for various reasons, neither method has found universal use. The boundary integral equation method (B.I.E.M.) can be more efficient and economical than the finite element method for 3-D problems involving large volume to surface area ratios. A B.I.E.M. program developed for calculating stresses and deformations in infinite elastic media has been used to determine the 3-D distributions of stress around T- and Y- junctions of circular tunnels in stress fields in which the ratios of horizontal to vertical stress are 0, 0.33, 1.0 and 2.0. For the T- junction considered,stress concentrations at the intersection are approximately twice the plane strain values. Except where horizontal field stresses are high, the stresses in both tunnels approximate the plane strain values at about two tunnel diameters from the point of intersection of the tunnel centre lines. The stresses in the two branches reach those for the corresponding plane strain, two parallel tunnel case two diameters from the point of intersection of thq inner tunnel walls. Very high stress concentra- tions occur in the pillar of rock between the inner tunnel walls near the point of intersection.

Jan 1, 1976

The lignites of Otago and Southland offer a potentially viable future feedstock for electricity production and possible conversion to petrochemicals or liquid fuels. The deposits, when combined, represent New ZealandÆs largest energy inventory; equivalent to about 74 000 petajoules û or approximately 35 times current natural gas reserves. Too often dismissed as irrelevant to New ZealandÆs energy future because of adverse perceptions of the value of low rank coals, the strategic value of this resource is in fact huge.   Investigations of the lignites extend back over more than 50 years (LFTB various, CRL Energy Ltd). Whilst it is not the intention of this paper to review this work, these investigations have confirmed the suitability of the resource as a feedstock for thermochemical conversion. Current estimates of South Island lignite mining costs are NZ$1.0 - 3.0/GJ (Natusch 2005), making these resources amongst the most competitively-priced energy sources available world-wide.   This value, however, will only be captured if we as a country begin to envisage a world-scale, long-life, integrated and multi-faceted development sized to supply world markets for chemicals or fuels. Recent CAE studies (Hooper and Riddolls, 2005) have emphasised the vital importance of ensuring that any exploitation of this resource maximises their value to New Zealand. The question that thus must be applied is how might large-scale utilisation of the resource fit with New ZealandÆs primary energy supply, and its potential for the export of high-value commodities.   From the broader perspective of New ZealandÆs total energy supply equation we need to examine other supply options beyond just the generation of electricity. We face in this country a situation wherein almost all our liquid fuels feedstock has to be imported with projections for an increasing dependence on imported oil. In addition NZ has a natural gas reserves-to-production ratio of less than ten years, and dwindling supply from local production. Without further natural gas discoveries, New Zealand has only a window of approximately five years before we will need to commit to imported LNG or some other thermal fuels option.   In other words, New Zealand is facing an uncertain future unless we can restore energy inventories to cover an adequate strategic reserve capacity as well as enable long-term investment in alternative sources. Within this context the South Island lignite resources are thus deserving of more attention. The extent of the resource is well known, its suitability for conversion reasonably assessed, and its magnitude such that there is little doubt that lignite conversion could adequately meet NZÆs future energy needs for many years out.   This paper reviews coal gasification technology, its application to lignite and the possibilities for future development of the lignite resources. It is argued that the strategic imperative for this country is to secure the lignite resources as a primary energy source for this country well into the latter half of this century.

Jan 1, 2005

By Kadoya H, Kondo Y

The Harima works were established in 1964 and started operation in 1966. The initial capacity was 3500 t/month of zinc and 1500 t/month of lead. Until 1990 a lot of improvements had been made and the capacity had increased 6600 t/month of zinc and 2200 t/month of lead without enlarging the furnace size. During the ten weeks campaign shut-down at the end of 1990, many modifications were carried out including the enlargement of the furnace shaft area. By these changes, the capacity has gone up to 7500 t/month of zinc and 2500 t/month of lead. In the sinter plant, the amount of SO2 rich gas supplied to the acid plant was raised to 50 000 Nm3/h from 43 500 Nm3/h. An electrostatic precipitator was adopted instead of a venturi scrubber to clean the rich gas. Oxygen enrichment has been continuously operated since August 1991, because the furnace capacity is more than the rate of sinter production. Utilisation of the sinter plant is high being as much as 97 per cent and the sulphur burning rate has been increased. In 1990 the furnace was replaced except the hearth. The shaft area was enlarged to 19.4 m2 from 15.3 m2. The target blast was designed to be more than 36 000 Nm3/h. Tuyere layout was changed to prevent slag breakouts. Due to these modifications shaft condition has been very good. Automatic launder jacket cleaners were installed to increase the lead cooling efficiency. Many other improvements and developments were also made. Coke powder injection from the tuyeres has been successfully operated since April 1987.

Jan 1, 1993

The Southern 'B' lode orebody at the New Broken Hill Consolidated Limited (NBHC) mine, Broken Hill, is a continuation of the zinc-rich orebody that was first mined in NBHC in the 1940's. Preliminary diamond drilling in the early 1960's resulted in sufficient information to delineate the southern extension of the presently mined 'B' lode orebody, In 1966 the sinking of a 7.3 metre diameter circular shaft (No. 3 Airway) was commenced to improve the ventilation of the southern extremities of existing workings and to provide a base to further explore and develop the orebody extension. Development on four levels was started in the early 1970'.s to link the existing workings with the new shaft. Three possible mining methods - sublevel open stoping, loader cut and fill with ramp access and sublevel caving - were investigated and the decision was made to mine the orebody by transverse sublevel longhole stopes and pillars with trackless equipment and ramp access between the major working levels.

Jan 1, 1977

By Pratt V

The economic, social and political environment of the mining and minerals industries is, as we all know, volatile and affected by constantly changing international forces. In addition, the outcome of negotiations with overseas customers is often uncertain for long periods, or even unfavourable. The industries and companies represented at this Conference are in an arena of high competition, risk taking and enormous capital outlay. They are industries which our overseas markets (rightly or wrongly) have criticised for industrial relations problems, unreliable production and uncertain deliveries. They are industries which have an urgent need to be cost effective and competitive at all levels of activity.A critical factor in the mining and metallurgy industries is, therefore, its human resources and its human resource management. It requires the full use of available talent, innovation, skills and qualifications within its workforce, and the appropriate management of these resources.

Jan 1, 1985