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BHP Billiton Olympic Dam produces copper cathode, uranium oxide concentrate and refined gold and silver. The operation is comprised of an underground Mine and Processing Plant and is located 560 km north of Adelaide. Water for the operation is sourced from two wellfields located 120 and 200 km from the operation in the Great Artesian Basin (GAB) at a rate of approximately 36 ML/d. Regulatory drawdown limits were established to protect regional GAB springs and are monitored at compliance bores. Drawdown was predicted in the æOlympic Dam Environmental Impact StudyÆ (Kinhill, 1997); however, pumping at higher volumes may have the potential for increased drawdown resulting in a trend towards the compliance limits. In recent years increased usage of water in the Concentrator and Hydromet sections of the Processing Plant has caused excess process water reporting to the Tailings Retention System (TRS). This has resulted in excessive amounts of liquor ponding on the tailings cells with a potential impact being contamination due to seepage of liquor to groundwater. Effective management of water by the operation is necessary to both reduce GAB drawdown and reduce the amount of liquor on the tailings cells. A site working-group has been established to investigate and pursue opportunities to improve the water use efficiency of the operation. This paper will highlight the issue of water management at Olympic Dam and discuss the various options currently being investigated to improve management of this important resource.

Jan 1, 2006

By H S. Pescini

A material aim when feeding ore blends into a processing facility is to reduce the variability of the properties of the feed material and to maximise the output, or throughput. At the Mining Geology conference of 2019, we presented a paper on optimal blending of coking coal given geological variability. The paper introduced integer-programming techniques to both maximise value and minimise feed or product variability. We discussed an application where the geological uncertainty is characterised using a normal distribution and exact solutions are readily obtainable using second order conic constraints in the mathematical model. Geological uncertainty is not always normally distributed, many properties will be characterised as log normal or some other skewed distribution. For example, copper can be some form of lognormal and Fe in Iron ore has a left skewed distribution. In this paper, we discuss some distribution free techniques to optimise the blends of plant feed or output given a quantified geological uncertainty. We discuss the sample average approximation method and how to set-up and apply the method to an ore blending problem. Evaluating a case study demonstrates that the explicit use of non-normally distributed geological uncertainty can better optimise feed into an ore processing facility and potentially add significant value. The approach is equally applicable to all mining and processing operations that blend material into a processing facility or directly into a saleable product.

Mar 22, 2022

By Creelman C. A

With the technique and data obtained in using the blast hole diamond drill for "Removal of Horizontal Level Pillar by means of Blast Hole diamond drilling at New Occidental Gold Mines N.L.," as described in Proceedings No. 133 of 1944, and a series of experiments using the blast hole diamond drill for stoping at New Occidental and New Cobar Mines over the past two years, a benching method of stoping has been developed using 2 inch coring diamond drill bits, resulting in efficient and economical ore extraction in these large ore bodies.During the experimental stages the standard H-inch blast hole was used, but the deflection of the holes, when using this solid bit, either decreased or increased the burden of the hole when it became necessary to drill holes deeper than 45 feet.It was found that after blasting vertical holes deeper than 45 feet, the ore did not break satisfactorily over the full length of the hole; on one occasion as much as 30 feet of bench was left at the bottom of the face, which had to be subsequently re-bored andblasted.As this defect was also noticeable to some extent even when the holes did not deflect, it was then obvious that the explosive had become less sensitive in the lower section of the hole. This was probably brought about by pressure being built up ahead of the detonation wave in H-inch hole, resulting in desensitization of the explosive.Another series of experiments were carried out, using 2-inch solid bits, which enabled a much greater burden on the holes, and the use of primacord for the full length of the hole.This gave satisfactory results regarding breaking and fragmentation, but was uneconomic owing to the excessive consumption of diamonds, and the slow rate of drilling.Further experiments were then carried out using 2-inch coring bits, (Fig. 1) which allowed holes of 90-feet to be bored with little deflection, and proved to be entirely satisfactory as to cost and speed of drilling.

Jan 1, 1945

By B Crompton, A Berghorst, G Knox

Laboratory-based dynamic testing provides an invaluable tool in the development of rock bolts for use in highly stressed, burst prone rock masses. While in most cases the laboratory-based test method is not a direct comparison to an in-situ installation due to the purely axial loading and variations in boundary conditions, it does provide an indication of the performance of the tendon system under dynamic loading. The aim of the research described in this paper is to provide an understanding of the effect of the impact velocity of the impact mass during a laboratory-based dynamic test. Understanding the effect of boundary conditions of laboratory testing will improve Geotechnical Engineers understanding of the performance in an in-situ installation. In addition, due to the limited number of dynamic test facilities, an understanding of the effects of the boundary conditions will aid in the comparison of test results for different products from different test facilities. The research was conducted using an impact tester, whereby a known mass is raised to a known height defining the kinetic energy and velocity of the mass at the point of impact with the test sample after release. The impact velocity was varied between 2.7 m/s and 5.4 m/s, while the kinetic energy of the mass was maintained at 11.5 kJ. While a correlation between the impact velocity and strain rate of the system tested was established, due to the properties of the system tested, the range of impact velocity on which dynamic tests were conducted was insufficient to result in a considerable change in the average impact load, the cumulative proximal displacement and the cumulative maximum energy absorbed over multiple impulse. However, due to the reduction in the momentum of the impact mass the impact duration and displacement during each impulse in the series of impulses reduced. CITATION:Knox, G, Berghorst, A and Crompton, B, 2018. The relationship between the magnitude of impact velocity per impulse and cumulative absorbed energy capacity of a rock bolt, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 160–169 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

Nickel laterite HPAL circuits are known for their complexity and have presented a large number of challenges to process engineers worldwide. An attitude commonly encountered by the author is that the beneficiation of the ore is relatively simple and that the difficult problems are in the realms of hydrometallurgy. On a number of occasions beneficiation has become the 'poor cousin' in the process design phase. Another common theme is that significant difficulties are encountered when beneficiation circuits are started up, invariably slowing the ramp-up of the remainder of the circuit. In some cases the problems impact on throughput for a long period and can require major rectification capital. The HPAL Ni laterite projects encountered each have their own set of beneficiation problems that need to be understood before an appropriate circuit can be designed. This paper will discuss a range of design issues including geometallurgy, ore characterisation through laboratory testing, the application of appropriate sizing and analytical procedures, the selection of appropriate unit operations, materials handling of plant feed and products and setting meaningful plant design requirements.

Jan 1, 2008

By M Szwec, P Corbett, P Sheffield

Centennial Coal operates the Springvale and Angus Place longwall mining operations that extract coal beneath the Newnes Plateau in the Western Coalfields of New South Wales. Springvale and Angus Place mines have a history of difficult geotechnical conditions, where mining operations are conducted with a thick coal roof interbedded with numerous claystone units and relatively high levels of vertical and horizontal stress. The combination of a relatively high stress regime and weak laminated roof has led to numerous roof failures throughout the mine’s life.Following underground roadway development, the mine roadways can experience high levels of roof movement that can lead to roof falls. Two-point wire extensometers (known as ‘telltales’) are extensively used for roadway condition monitoring. Along with extensive other geological and geotechnical data, monitoring data is used to prepare geotechnical hazard plans that are then used in all areas of mine planning to allow for safe and efficient operation of the mine.This paper presents a case study of the use of a new software tool that allows time-scaled visualisation of monitoring data in the context of longwall and development face positions and installed strata support. It can be used in the identification of geotechnical risk factors and their interaction with the mining process to allow a more robust geotechnical hazard process. It interfaces directly with the widely used ExtoChart database software and can be presented with scale drawings of geological/geotechnical data to allow correlation of monitoring data with risk factors.CITATION:Corbett, P, Sheffield, P and Szwec, M, 2014. A new tool for extensometer data analysis and improved understanding of geotechnical risk factors, in Proceedings AusRock 2014: Third Australasian Ground Control in Mining Conference , pp 217–232 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 5, 2014

By I C. Carrasco, V G. Mirada, J F. D Domingo, F S. Leite

This paper investigates ore loss and dilution produced by rock movement in bench blasting. A total of four blasts, in which electronic detonators have been used, were monitored in Cobre Las Cruces Open Pit Mine (FQM Ltd) in Gerena (Sevilla, Spain). Rock motion has been obtained in 137 points of the free face and 6 points inside the block. Multiple liner regression has been applied to predict rock motion and estimate ore dilution as a function of hole coordinates, rock contacts, powder factor/hole, amount of explosive per blasthole, burden, spacing, hole length, stemming length, explosive energy, hole row, subdrilling, blast sequence, stiffness ratio, blasthole distance to free face, rock density and Young’s modulus. The model is statistically significant with a determination coefficient over 0.8. Rock motion measurements in 31 more points have been used to validate the model and to show a reduction in ore dilution.CITATION:Domingo, J F D, Leite, F S, Mirada, V G and Carrasco, I C, 2015. Dilution, ore grade and blast movement calculation model, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 71–78 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By P J. Chapman

Recent catastrophic failures of tailings storage facilities (TSFs) continue to highlight the need for a detailed understanding of failure modes and the defences that are in place within the design to mitigate the failure modes. However, in many cases the instrumentation devices and monitoring approaches selected during operation do not link back to the design defences and inappropriate or redundant instrumentation is often installed. Candidate instruments and monitoring approaches (critical controls) are presented and matched to critical failure modes, with clear linkages to the intent of each instrument/monitoring approach. Commentary on where complementary measurements with independent technology types are recommended to provide greater assurance of trends to the design engineer or TSF owner is also presented. In response to the recent TSF failures, an increasing focus is being placed on instrumentation and monitoring relating to slope instability and safeguarding against the potential for static liquefaction, using a critical state soil mechanics framework. To support selection of appropriate zoning within a stability model, three complementary devices are recommended – vibrating wire piezometers to provide an indication of pore pressure and allow the phreatic surface to be inferred, tensiometers to confirm the presence of unsaturated conditions (and hence low potential for undrained failure) and moisture content sensors to provide an indication of void ratio, and hence state parameter, in saturated zones. The combination of VWPs and tensiometers will provide complementary measurements in the saturated and unsaturated zones, and also provide information on the extent of any near-saturated zones, provided that sufficient instruments are installed.

Jul 1, 2021

Open pit mining at Bingham Canyon commenced in 1904 and about two billion tonnes of ore have been extracted since then. There remain just under one billion tonnes of ore in the pit and this will be mined by 2015. Because the porphyry system extends below the ultimate pit limits, Kennecott Utah Copper Corporation initiated a study in 1997 to determine the viability of underground mining. The study was initiated at that time because the results were to be used to compare the economics of a further pit expansion to that of an underground mine. If a pit expansion was viable then overburden removal would have to start in the near future. This paper is a summary of the work that was carried out for the underground study. Block cave mining was determined to be the most appropriate method to mine the deposit. Two block cave areas have been planned with a total production rate of 72 500 tpd. Additional sources of ore will come from skarn orebodies that are associated with the porphyry deposit. The geology of the deposits is described together with the geotechnical techniques that were used to characterise the rock in the absence of more detailed data that are normally available. A general description of the mine plan is given along with the reasons for selecting the particular methodology used. The study was completed at the end of 1998 and demonstrated the viability of underground mining at Bingham Canyon. Funds have been approved for further drilling and technical studies, which will lead up to a feasibility study in 2004/5.

Jan 1, 2000

By G N. Phillips

Global annual gold production has been remarkably constant over the last decade at ~2500 t/a3. The major producers today are China (annual production over 350 t and rising), Australia (250 t), USA (240 t), Russia (200 t) and South Africa (190 t and falling); these figures are representative of recent years but approximate. Demand from jewellery, investment, industrial and dental uses has been in excess of mine production for the last decade, by approximately 1000 t/a.Ongoing exploration is the life-blood for gold mining, and more so than for many other metals. Explorers consider sovereign risk, security, labour costs, commercial terms, land availability and inherent geologic prospectivity to determine where to focus their activities. Exploration success over the last decade has been shared by established major producers like Australia, Canada and USA, several countries in West Africa, and many parts of the Pacific margin. Although gold-only deposits have dominated world gold production historically, the gold-plus types, especially copper-gold, are becoming more important.Australia experienced a major period of exploration success from 1979 that has led to annual gold production rising from 17 t in 1979 to 250 - 300 t/a recently. Today, Australia has economic demonstrated resources of over 9000 t Au, which has risen progressively from 220 t in 1979. The majority of production has come from gold-only deposits in the Yilgarn craton in the southern half of Western Australia, but there are a few large Proterozoic deposits in central Australia including Olympic Dam, Telfer and Starra, which are mostly gold-plus types with copper and gold. Large goldfields in the younger host rocks of eastern Australia include the gold-only deposits of Victoria and North Queensland, and gold-plus deposits of New South Wales.CITATION:Phillips, G N, 2013. Australian and global setting for gold in 2013, in Proceedings World Gold 2013 , pp 15-22 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013

By B W. Atkinson, J Docherty

An Australian Working Group has been compiling a guide for operators to use in order to improve the efficacy of their routine and non-routine sampling tasks within their coal preparation process. The guide focuses on unit operations within the process, and not conveyor belts or transfer stations that are adequately covered by the existing Standard ISO 13909.Representative sampling of unit operations is generally an after-thought following plant construction, whether that be a result of oversight, or more commonly, capital cost minimisation during initial construction. Minimising capital cost by avoiding adequate sampling systems is generally a poor decision, since the result is a process that cannot be adequately measured, meaning that it cannot be adequately managed and optimised.The use of the term ‘guide’ rather than ‘standard’ is intentional. Practical means of sampling unit operations within existing plants generally relies on a compromise between sample representativity (minimisation of bias) and sampling feasibility. The guide does not endorse poor sampling practice, but instead offers practical guidance to overcome, to the extent reasonably possible, real-life limitations.When it comes to sampling any particular process, it is almost always necessary to make a professional judgement as to whether or not the proposed samples are likely to be sufficiently representative such that the sampling will be worthwhile. Such judgement needs to be exercised by personnel who understand the principles of sampling, as they apply to the stream being sampled. There are many circumstances where available methods for sampling of a particular stream will be meaningless, and in such circumstances analytical data from that sampling will only serve to propagate poor decision-making.This paper outlines the approach for the guide, the areas it covers and the approaches adopted for different types of unit operations and associated equipment configurations.CITATION:Atkinson, B W and Docherty, J, 2017. A practical guide to sampling in coal preparation plants, in Proceedings Eighth World Conference on Sampling and Blending , pp 241–246 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2017

By D W. Wall, J P. Vagenas

Questions of sampling precision and accuracy affect every process plant in differing ways and to differing extents. To address these perceived issues, chemists, metallurgists and engineers are often asked to check their results to confirm their reliability. However, it is not possible to achieve the level of detail required to check every sample, especially if the event has already occurred. No operation has the resources to frequently sample every point in the process for the purpose of verification.One way to address this scenario is by generating a sufficiently detailed comparative data set from a dynamic mass balance simulation of the processing plant. This mass balance uses both sample and plant instrument data to model every flow, while also taking inventory changes and process chemistry into account.The mass balance data can be used to identify any bias present in plant sampling and analysis. It can also be used to verify any unexpected assay results by taking a holistic view of the process. Improving the precision and accuracy of sampling data helps support operational decision-?making and understanding. It also improves the accuracy of the dynamic mass balance used in this process, facilitating further improvements.CITATION:Vagenas, J P and Wall, D W, 2017. The benefits of automated metallurgical accounting for plant sampling, in Proceedings Eighth World Conference on Sampling and Blending , pp 203–210 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2017

By A Marton

Concerns with crest losses have seen changes to a number of design and operating parameters to retain the integrity of the wall and the berms of the Kidston Gold MinesÆ Eldridge Pit. Until trials began in 1999 most final wall designs utilised large diameter blastholes up to the presplit row to achieve the wall angle. The berm crests were being lost, wall support requirements were increasing and affecting mining schedules and several fault areas were putting pit access at risk. Standard designs had been determined during the early mining years and while periodic modifications had been made there was a need for major blast pattern revision. This paper traces the development of wall control blasting programs over the years, relating designs to the geology and the evidence in the walls with the theories of the day and includes the recent trials that were conducted in what is a most unforgiving geology. Trials and improvements to ensure the walls were safe were conducted even as the mine neared its economic life. Instrumented monitoring of trials blasts has greatly increased the understanding of the damage criteria and enabled rapid revision of designs. Recent changes included reduction in presplit energy, charge redistribution and use of smaller blast holes for the trim patterns. Determination of a minimum proximity-to-final wall limit for production blastholes, establishment of the correct standoff to berm crests and walls and selective use of presplitting techniques for protecting crests have resulted in cost effectiveness, improved bench crest retention and an overall reduction in visible wall damage.

Jan 1, 2001

By Y Wang

Inadequate goaf gas control may lead to gas exceedances in the ventilation return air, which will trip power in longwalls and cause production stoppage. To effectively control gas emissions, Australian mines heavily rely on drilling vertical goaf holes to capture coalmine methane before it migrates to mine openings. The wide application of goaf gas drainage may change the goaf pressure distribution, lead to more ventilation air leakage, and consequently increase gas explosion and spontaneous combustion risks. This paper focuses on analysing goaf gas drainage production data from a case study longwall mine in Australia. The correlation between suction pressure, total flow rate, gas flow rate and air leakage rate were analysed in detail as the face-to-hole distance changes. Suction pressure is found to be positively correlated with total flow rate, and their correlation also varies at different goaf locations due to various goaf compaction and flow resistance. This paper also proposed a theoretical method to calculate the goaf resistance of ventilation air leakage pathways. Based on several basic assumptions and different goaf drainage scenarios, the leakage airflow pathway resistance in the goaf at different face-to-hole distances can be calculated. The model results suggested that the resistance of air leakage pathways from the working face to individual goaf holes increased with the face-to-hole distance. The resistance results can also be used to calculate the corresponding permeability at the same position, which can be used as input data for CFD or Ventsim models.

Oct 12, 2022

By B Crowdey, J Wall

The particular underground operation relating to this paper is a gold producer and has an orebody predominately consisting of rhyolite and marcasites pyrite. The ore has been identified as being reactive to nitrates.Operations currently utilise two twin boom Tamrock™ Super-Drill Jumbos for development mining. Capital (waste) headings have a nominal size of 7 m wide × 6 m high, while ore headings are designed at 5 m wide × 5 m high, with the lower reaches of the mine narrowing down to a 4.5 m × 5 m design. Production drilling is currently undertaken with an Atlas-Copco L6 Solo. Due to the reactivity of the ore, inhibited emulsion phase bulk explosives are used for all blasting activities, with electronic detonators utilised in the production sequence.Sublevel shrinkage stoping is utilised when the orebody is stronger than the host rock. After blasting the ore, and during extraction, the surrounding host rock cannot maintain integrity and it ‘collapses’ on top of the ore. Gravity forces this collapsing rock to follow the blasted ore through the drawpoints.The highest risk involved with sublevel shrinkage stoping is the necessity for charge crews to work upon ‘rills’ at the drawpoints, which are occasionally unstable, to charge the next blast ring to be fired. To overcome this issue a loading technique called ‘precharging’ is used whereby two to three rings are charged, but only the first ring is fired. This technique can minimise risks by taking the charge crew away from the rill and brow.As this mine contains hot and reactive ground, strict guidelines from the Australian Explosives Industry and Safety Group Inc (AEISG) are adhered to. These guidelines relate to the sleep time for the precharged rings.This paper describes the issues faced with precharging in hot and reactive ground, the monitoring required and the systems and procedures that were put in place to enable a successful outcome.CITATION:Crowdey, B and Wall, J, 2014. Precharging sublevel shrinkage production rings in hot and reactive ground with electronic blasting systems, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 283–288 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By R E. McIvor, K M. Bartholomew

Classification system efficiency (CSE) is a simple, powerful metric that provides a comprehensive measure of classification system performance. CSE directly quantifies the amount of coarse solids in the mill, and as a result, also directly quantifies the percentage of mill power being used effectively. In this manner, it is an effective, practical grinding energy efficiency metric.Three independent variables drive CSE: circulating load ratio, sharpness of classification, and classifier bypass. Due to practical downstream constrains on per cent solids, improvements realised by increasing circulating load reach diminishing returns as sharpness of classification and bypass deteriorate in the most commonly used classification device, the hydrocyclone. This combination of effects leads to the use of CSE, and its relationship to operational economics, as an objective circuit optimisation criterion.Its simplicity in measurement makes CSE a useful tool for the plant metallurgist, and it is not subject to a great amount of experimental error. CSE provides the plant operator a fast, simple, and effective means to evaluate the performance of the plant classification system. Improvements in CSE relate directly to circuit throughput and energy efficiency. This paper, through description and plant case studies demonstrates the power and utility of this emerging grinding system optimisation metric.CITATION:Bartholomew, K M and McIvor, R E, 2013. Classification system efficiency - a comprehensive closed-circuit grinding optimisation metric, in Proceedings World Gold 2013 , pp 389-394 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013

The Karangahake gold-silver deposit is located 11 km west of Waihi in the southern part of the Hauraki Goldfield. Historically it was the second largest bullion producer in the goldfield with a past production (1882 - 1944) of 126.712 tonnes of gold-silver bullion (Au:Ag ratio of 1:3.3) from 1 142 315 tonnes of ore grading 27.4 g/tonne gold equivalents.

Jan 1, 1987

By D P. Blair

The theme of this conference is blasting productivity. This article only considers the technology aspect of this theme, and in particular, the collection and analysis of experimental data. In many cases, field trials are set up to collect data on variables such as fragmentation and vibration with the aim of comparing different explosive products and/or blast designs. An important aspect of such trials is the collection of sufficient data to draw sound conclusions. In this regard, there is no doubt that the scatter in real data is the root cause of all problems of obtaining misleading results due to an insufficient number of trials. This article shows that the typical one-off trial has a high probability of producing misleading information, which could be considered to be worse than having no information at all. Assuming that sufficient data has been collected, this article also highlights the methods that can be used to make the most of that data. These methods include the application of appropriate statistics that are dependent on the nature of the scatter in the data, and the use of some of the more recent filtering and spectral approaches that can be used to reduce the noise in measured blast signals and increase the resolution of their spectral characteristics.

Nov 8, 2011

By D Connelly

High purity alumina (HPA) is a processed, premium non-metallurgical alumina product characterised by its purity level ie 99.99 per cent (4N) or 99.999 per cent (5N). The market price, application, and performance of HPA varies widely according to its degree of purity. High purity alumina is a form of aluminium oxide (Al2O3) with growing use in LED lighting and coating on the ceramic cathode separator sheets in lithium ion batteries. This paper describes a study on producing high purity alumina from feldspar to produce HPA and potash for use as a fertiliser. The project was positive for a number of reasons. High potassium extraction (>99 per cent) was achieved in the alkali leach using sodium hydroxide as the leaching agent and full dissolution of the alkali leach residue was achieved using hydrochloric and sulfuric acid. The potential to produce a high purity alumina product via a complex but promising emerging process was assessed. The potential to extract potassium as potassium sulfate to add value to the project was also demonstrated with the test work.

Sep 1, 2024

By C D. Pocock

Maintaining catch berm capacity in open pit mines through controlled blasting prevents production delays associated with managing risks associated with wall instability. A method has been developed to quantify the drill hole deviation of presplit holes at Century Mine. Different controlled blasting techniques are used to control wall quality and decrease the risk. The performance of presplit drilling, in terms of deviation from design has been investigated. A methodology has been developed to quantify the drill hole deviation in three dimensions. This methodology was converted into a series of algorithms capable of handling an extensive range of data. The methodology was applied to a case study at Century Mine to evaluate the performance of drilling on two presplit blasts. Overall, 67 drill holes were investigated in the case study. The drill hole deviation was quantified in terms of three parameters: skewness, linearity and drill hole deviation parameter. Methods for quantifying each parameter were developed. Skewness, a measure of the amount of curvature and linearity, was defined as the measure of the consistency of the gradient of the drill hole. Drill hole deviation parameter is defined as the average measured deviation of the surveyed drill hole position from the design position. Acceptable drill hole deviation tolerances for each of the defined parameters were developed by analysing drill hole deviation parameters with respect to implications to catch berm capacity. Verification of the methodology was confirmed through the comparing of the completed drill hole deviation results using both visual inspection and cross analysis for each of the three methods. Holes with an R squared value of less than 0.97 for skewed deviation were classified as having skewed deviation, while holes with a difference in gradient of 3.5 were classified as having linear deviation. The results show that 76 per cent of the drill holes had deviation. Of this, 19 per cent were skewed, 57 per cent had linearity and 12 per cent contained both skewed and linearity deviation.

Nov 8, 2011