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An efficient and portable mathematical model has been developed for simulating heat transfer in six-in-line slag-resistance-heating electric furnaces for smelting sulphide ores to produce base metals and platinum group metals. This model is a steady-state one relating furnace conditions and performances to various control and input parameters. Some transient effects occurring in electric furnaces are neglected for computation efficiency. This article describes the model development and modelling results. The present model is capable of predicting: (i) temperatures at various locations in a six-in-line furnace, such as slag bath, matte bath, solid charge, freeboard space, freeze lining, cooling water, and air gaps between solid material components, etc. (ii) freeze lining thickness, (iii)smelting rate, and (iv) heat loss rate, etc. A typical feature of the model is that it is easily portable to different application platforms and sufficiently efficient with execution times less than a few seconds. Therefore, it is possible to apply the model for online prediction and control of heat transfer and freeze lining thickness in industrial electric furnaces.

Jan 1, 2011

By Pattinaja MA

Mathematical modelling was carried out to determine the probable convergence and vertical stress patterns around a retreating longwall panel in an underground coal mine in New South Wales. An 'Electrical Resistance Analogue' and a program called 'THREED' based on the displacement discontinuity method were used for the modelling. Parts of the obtained results are presented here and compared with some measurements of convergence at the site. The modelled and measured data agree better qualitatively, than quantitatively.

Jan 1, 1986

By P. A. Bahri, D. G. Dixon, T. M. Woodward

A comprehensive mathematical model of the Sherritt-Gordon ammoniacal pressure leaching process at K winana Nickel Refinery has been constructed. The process is comprised of six autoclaves functioning as a three-stage leach circuit, operating over the temperature and pressure ranges of 85-95°C and 750-1000 kPa, respectively. The process chemistry forms a complex series-parallel reaction network, characterised by: (1) the gas-liquid mass transfer and reduction of oxygen, (2) the oxidation of eight distinct minerals, (3) the oxidation of sulphur, and successive reactions involving various intermediately oxidised sulphur species, and (4) the oxidation of Fe(II) to Fe(III), and subsequent precipitation as a hydrated ferric oxide. Material balances were developed for species in all three phases, and an energy balance to account for all reactions and phase conversions. The multiple convolution integral (MCI) statistical reactor model was employed for the scale-up of particle kinetics. The model has been validated against plant data, and simulation results are in agreement with plant performance.

Jan 1, 2005

By Vladimir Panjkovic, Ostrovski. Oieg, Campbell Cripps Clark, John Truelove

"Effects of properties of raw materials and blast on the performance of an iron bath reactor have been examined by a mathematical model. The model consists of three modules: 1) chemical reactions in the gas phase and heat transfer from gas to droplets ejected from the bath; 2) heat losses from the vessel and radiative heat transfer from the freeboard to the bath; and 3) heat and mass balances for the bath. The following parameters of the process are calculated: coal, ore, flux and oxygen consumption, post-combustion ratio and heat-transfer efficiency, temperature and composition of the off-gas, and slag and metal chemistry.IntroductionOne of the anticipated advantages of smelting reduction over conventional ironmaking is the use of lower-grade raw materials. In this work, the effect of-properties of raw materials on the performance of an iron-bath reactor is examined with a numerical model. The model is also used to assess the effect of properties of the blast (flowrate, temperature and oxygen enrichment) and ejected droplets (generation rate, size and ratio of total weights of slag and iron droplets) on the key parameters of the process. In the process coal, oxygen, ore (which may be partly prereduced) and fluxes for specified slag composition are charged to the bath. Slag and iron droplets are ejected from the bath into the gas phase, where they are heated before returning to the bath. Iron droplets undergo decarburisation and reoxidation with CO2 and H2O. Dust is also generated from the bath, and oxygen or preheated air for post combustion is injected from the top of the vessel.The model consists of three modules: 1) reactions in the gas phase and heat transfer from gas to droplets ejected from the bath; 2) heat losses from the vessel and radiative heat transfer from the freeboard to the bath; and 3) heat and mass balances for the bath. The structure of the model and the data exchange between the modules are shown in Fig. 1. The following parameters of the process are calculated: coal, ore, flux and oxygen consumption, postcombustion ratio (PCR), heat transfer efficiency (HTE), off-gas temperature and composition, slag and metal chemistry, and rates of decarburisation and reoxidation of the iron droplets."

Jan 1, 1996

By Isobe K, Muchi I, Hsieh YS

Theoretical and experimental studies have been made-to clarify the physical and chemical behavior of the tuyere combustion zone of blast furnace. Gas streamlines around this zone were visualized by the use of the two-dimensional equip- ment. Close agreement between the observed stream- lines and the calculated ones is found. The three- dimensional experimental studies have revealed that there exists a so-called dry zone in the neighbor- hood of the raceway and the trickling liquid be- haves approximately like a continuum. A kinetic model is develdped to describe the behavior of coke combustion in the tuyere zone. The model is based on the reaction kinetics, the differential balances of heat and mass, and the mechanics of gas flows. Regarding-the-axial distribution of gas composition in the tuyere zone, good agreement is obtained be- tween the model prediction and the data observed in a working furnace by the other investigators. Cal- culated results-show that the maximum consumption of coke occurs at the intermediate region of the raceway. For the blowing of oxygen-enriched or high -humidity blast, the position at which the maximum consumption of coke occurs is shifted toward the tuyere nose. Similar effect is found when the par- ticle diameter of coke in the tuyere zone is de- creased. The solid flow pattern around the tuyere combustion zone was visualized by the use of the two-dimensional equipment in which the sublimation of dry ice particles in front of the tuyere was

Jan 1, 1981

By A. L. Mular

Some specific cases of the collection of data from industrial grinding and classifying operations, and the development of models from the data so obtained, will be described. The problems encountered, the effort involved, and the value of the derived models in improving the plant efficiencies will be discussed and documented where possible.

Jan 1, 1971

By Dan G. Cojocariu, E. Ene-Danalache, I. Huber-Panu

A general model which represents not only batch but also continuous flotation in multi-cell machines is presented. This model considers both the size distribution and the distribution of flotabilities in a size class for valuable mineral particles in the pulp. The general expression is shown to reduce to several previously reported models. In addition, a number of new models are found to arise for different specific assumptions concerning the flotability distribution and/or the size distribution of the particles undergoing flotation. Several of these new models are tested experimentally using laboratory-scale batch and pilot-scale continuous data. Model parameters are estimated from this data and a comparison of calculated recoveries and those experimentally determined are presented. Finally, the influence of the feed rate, aeration rate and valuable mineral content of the feed on the recovery is examined.

Jan 1, 1976

By R. Venkataramani

Crushed stone operations have become more difficult due to rising costs of labor and equipment, with the consequent reduction in profit margins. Competition from other operators and the encroachment of vertically integrated companies place severe restrictions on available markets. The seasonal demand for stone precludes any definite production level, and changes in productivity have to be accommodated. These influence profitability, and therefore production plans must be continuously adjusted to keep pace with rapidly changing conditions. The basic problem of stone companies seeking to maximize profits is one of balancing stone production costs, costs of inventory holding and stock-piling, marketing costs and any other charges accruing over time against the selling price of stone. Because of high peak demands, slack period production must be stored and then distributed to supplement peak period production. For a multi-plant operation, there is a need to evaluate at the same time the interaction between the plants and the various marketing sectors. In a multi-plant operation with different production capacities and production costs, stone producers are interested in the answers to the following questions: 1. How much stone should be produced at each plant in each time period? 2. What should be the optimum stone shipment from each plant to each market? 3. What are the effects on net returns and shipping plans, if production capacities are altered at the plants? These questions have many different answers that, will satisfy demands. From the point of production scheduling, the minimum production cost may be the objective. From the inventory standpoint, the best solution will be to synchronize production with sales. However, when the transportation of stone is also considered, the desired end is to achieve that delivered cost which maximizes the company f s net return. Besides, only an extremely naive analysis ,wouId attempt to

Jan 1, 1971

By Saul Garcia Hernandez

The research work objective is to study a new ladle shroud internal design to reduce the slag emulsification in the tundish during ladle transient periods. A 3D mathematical model of the tundish was created to solve the fundamental equations, a turbulence model, and a multiphase model. The results show that using a conventional ladle shroud, the fluid is delivered with an excessive amount of kinetic energy being dissipated inside the tundish bath, generating strong mixing flow patterns and entrapping a massive amount of slag. In the other hand, the proposed ladle shroud dissipates the kinetic energy before the fluid enters the tundish promoting less intense mixing patterns; thus, the amount of slag emulsification is reduced significantly. Consequently, if the internal design of the ladle shroud is able to dissipate the kinetic energy before the fluid enters the tundish, it shall be possible to reduce considerable the slag emulsification and the slag opening area.

Aug 17, 2017

By S. Sarac

In this study, graphical and mathematical methods used to determine operating points in single or multiple fan ventilation systems are summarized. The problem is mathematically formulized for parallel connected fans in which analytical solutions are not possible. For n number of parallel connected fans, a non-linear equation system that consists of n number of non-linear equation sis derived. The Newton-Raphson method, which is a numerical analysis method, is adapted to solve this non-linear equation system and the solution steps are given as an algorithm. A computer program based on this algorithm is written and its applicability is proved.

Jan 1, 2000

By A. E. Hilder

Introductory The following pages present the results of field investigations, laboratory and commercial tests, drilling, surveying, and preliminary development of the refractory clays which have been discovered recently in the township of Kipling, at the foot of Long rapids on the west side of the Mattagami river about 57 miles north of Kapuskasing and approximately 7 miles north of Smoky Falls, the terminus of the privately owned railway operated by the Spruce Falls Power and Paper Company, Kapuskasing (see Figure 1). Investigations on the clays of northern Ontario have been carried on over a period of seven years by officers of the Ontario Department of Mines, and I am indebted to the information published by that Department in reports by W. S. Dyer and A. R. Crozier, as well as?by Professor R. J. Montgomery, of the Ceramic Department, University of Toronto. It was, in large measure, the data given in the 37th, 38th, and 42nd Annual Reports of the Ontario Department of Mines that led me to make a personal investigation of the Mattagami River clays, with the result that we now have a commercial undertaking that will develop eventually into an industry of considerable importance in supplying the central manufacturing districts of Canada with refractories that heretofore have been largely imported from the United States. An extract from the 42nd Annual Report of the Department, Part III, 1933, page 75, reads as follows: "As for the refractory clay, the deposits are undoubtedly very large, but what is urgently needed is a deposit near the railway under sufficiently shallow overburden for recovery by stripping'. Having previously made an investigation of the clays adjacent to the Temiskaming and Northern Ontario railway, and these clays proving unsuitable, I turned my attention to the outcrops of refractory clays on the Mattagami river, first reported by Keele in 1919. This area offered possibilities within the economically permissible distance from rail transportation.

Jan 1, 1935

Recovery of tin and copper from concen trates containing cassiterite, stannite and chalcopyrite was studied in laboratory-scale tests designed to simulate continuous operation of the matte fuming process in which tin is recovered as SnS fume while copper accumulates in the matte. Charges made up of 300g of mixtures of the concentrate and siliceous flux were added at a uniform rate in pelletized form to crucibles containing 300g of matte and 50g of slag (both of which were recovered from previous tests), while blowing through the matte with air or oxygen-nitrogen mixtures at a rate of 3 to 5 1/min. Accurate material balances enabled the effects of process operating conditions on tin recovery and sulfur utilization to be deter- mined. When operated correctly, the matte fuming process gave tin recoveries exceeding 90 percent in spite of wide variations in the copper content of the matte and in the operat- ing conditions. Deterioration of tin recovery was associated with undersaturation of the slag with respect to silica, and with high partial pressure of sulfur dioxide (greater than approx- imately 0.2) brought about by oxygen enrichment of the blowing air. The utilization of sulfur varied widely in low grade mattes. Copper in excess of approx- imately 10% promoted formation of SO2 by the commonly accepted reaction: FeS + 1-120 2= FeO + S02 while in copper-free and lower grade mattes a substantial fraction of the oxygen was consumed in the alternative reaction FeS + i 02 = FeO + z S2.

Jan 1, 1980

By Eric C. Halpin

"1. Introduction and Backgrounda. In its fundamental form, risk is simply the likelihood of an adverse impact. Expanding on this theme, the Corps of Engineers has broadly defined risk associated with infrastructure as a factor of three distinct elements: a probabilistic assessment of a hazard such as a flood or earthquake; a probabilistic assessment of how the infrastructure performs in the face of this hazard; and a measure of the likely consequences – loss of life, economic damage, environmental impacts – associated with the resulting inundation scenarios.b. The lessons and observations described herein relate to the Corps of Engineers experiences in implementing risk concepts for infrastructure management during the period 2005 through 2013. The Corps of Engineers is a large, decentrally operated and centrally led federal agency that plans, designs, constructs, operates, maintains, and/or manages a portfolio of over 3,000 built dam and levee systems within the United States. The infrastructure in this portfolio is 55 years old on average, and although much of the structures provide for flood risk management benefits, many of the dams deliver multi-purpose benefits. Scale matters – where scale and uncertainty are large, risk concepts provide great opportunity and arrangement for application.c. The amount of organizational change that is demanded for the implementation of risk concepts is considerable and should not be under estimated, but need not be feared if the path others have taken is considered. This paper is, in many ways, an opportunity to step back from the last seven years of implementing risk informed management, check our own azimuth and maturation, and share the purpose and energy behind our own continuing actions. Hopefully, this azimuth check is also beneficial to those that are either considering the use of risk or are in the process of implementing risk concepts.d. Success at managing the massive change related to incorporation of risk concepts will directly affect success of the infrastructure program. The three key organizational elements to consider –human systems such as governance and competencies; policies that outline the who, when, and why of decisions and management; and the processes and technology that outline how work is to be performed - are changed fundamentally. From senior leaders to project team members and all of the people in between, there is a need to discuss, confront, and vet the change and the role people and organizations have in the process. For the Corps of Engineers, with a history of over two centuries, deeply embedded beliefs about infrastructure aren’t transformed over night - the change process is decadal in scope. Given our self imposed “project execution” mantra and corporate objective, risk seemed threatening at first to the cost and schedule of delivering infrastructure projects. In reality, risk concepts have been responsible for savings and cost avoidance of nearly $6 billion over the last three years alone while at the same time justifying increased investment by several fold for the highest risk needs."

Jan 1, 2014

"INTRODUCTION Twin Metals Minnesota (TMM) (operating as a joint venture between Duluth Metals and Antofagasta PLC) assumed control of the property in 2010. Amec became involved with the project in 2011 and in 2012 incorporated additional drilling in order to update historic resource estimates. The Maturi geological model and resource estimate was updated in 2014 to incorporate additional information obtained during 2013 and 2014 drill campaigns. In 2017, Amec was purchased by the Wood Group and is now known as Wood. Geology The Twin Metals Minnesota (TMM) Maturi, Maturi Southwest, Birch Lake, and Spruce Road properties are located in northeastern Minnesota (Figure 1). This talk discussed the Maturi deposit only, although all four deposits share similarities. The following geology is summarized from Peterson and Severson (2002).Maturi lies within the Superior Province of the Canadian Shield which is exposed in central and northeastern Minnesota and extends northward into Ontario. In eastern Minnesota, the Midcontinent Rift System developed during crustal scale extension in the Mesoproterozoic Era. The rift system is traceable, as exposures of mantle-derived tholeiitic to subalkaline mafic lava flows, intrusives, and rift-filling fluvial sedimentary rocks, and in the subsurface as a gravity anomaly (high), from the eastern end of Lake Superior, arcing west across the lake basin, and extending south-southwest to northeastern Kansas. Intrusion of the Duluth Complex, circa 1,108 - 1,099 Ma, was related to rifting and is co-genetic with the North Shore Volcanic Group mafic volcanic rocks that form the hanging wall of the Duluth Complex to the east. Cu-Ni-precious metals (PM) mineralization including Pt, Pd, Au and Ag within the Duluth Complex is hosted by mafic to ultramafic intrusive rocks that are part of the overall intrusive complex. The Duluth Complex is defined as a more or less continuous mass of mafic to felsic plutonic rocks that extends northeastward for more than 170 miles (275 km) in an arcuate fashion from Duluth nearly to Grand Portage in Minnesota. It is bounded by a footwall of Paleoproterozoic sedimentary rocks and Archean granite-greenstone terranes (Peterson and Severson, 2002), and a hanging wall largely of comagmatic anorthosite, rift related flood basalts, and hypabyssal intrusions of the Beaver Bay Complex. Mineralization is hosted in the basal mineralized zone (BMZ) of the South Kawishiwi Intrusion (SKI). The SKI magmas intruded sub-horizontally between hangingwall Anorthositic Series (AN) rocks and footwall granitic rocks of the Neoarchean Giants Range batholith (GRB). The Maturi local geology and the deposit footprint (shaded orange) is shown in Figure 2."

Jan 1, 2018

Kroemer(1-3) and Kroemer et al.(4) define the isoinertial technique of strength assessment as one in which mass properties of an object are held constant, as in lifting a given weight over a predetermined distance. Several strength assessment procedures possess this characteristic. Most commonly associated with the term is a specific test developed to provide a relatively quick assessment of a subject's maximal lifting capacity using a modified weight-lifting device.(1,5) The classic psychophysical methodology of assessing maximum acceptable weights of lift is also considered an isoinertial technique under this definition.(6) While the definition provided by Kroemer(1) and Kroemer et al.(4) has been most widely accepted in the literature, some have applied the term "isoinertial" to techniques that differ somewhat from the preceding definition, such as in a description of the Isotechnologies B-200 strength testing device.(7) Rather than lifting a constant mass the B-200 applies a constant force against which the subject performs an exertion. The isoinertial tests described here apply to situations in which the mass to be moved by a musculoskeletal effort is set to a constant. Is Isoinertial Testing Psychophysical or Is Psychophysical Testing Isoinertial? As various types of strength tests have evolved over the past few decades, some unfortunate developments in terminology have arisen to describe and classify different strength assessment procedures. This is particularly evident in sorting out various tests that have been labeled "isoinertial." One example was cited earlier. Another problem that has evolved is that the term "isoinertial strength" has two connotations. The first is the conceptual definition: isoinertial strength tests include any strength test in which a constant mass is handled. In practice, however, the term is often used to denote a specific strength test in which subjects' maximal lifting capacity is determined using a machine and a constant mass is lifted.(1,5) Partially as a result of this dual connotation, the literature contains references to both "isoinertial strength test" as a psychophysical variant(8) and the psychophysical method as an "isoinertial strength test" (4,9) To lay the framework for the next two chapters, the authors will briefly discuss some operational definitions of tests of isoinertial and psychophysical strength. In stating that the isoinertial strength test is a variant of the psychophysical method, Ayoub and Mital(8) refer to the specific strength test developed by Kroemer(1) and McDaniel et al.(5) Clearly, this isoinertial protocol has many similarities to the psychophysical method: both are dynamic; weight is adjusted in both; and both measure the load a subject is willing to endure under specified circumstances. However, while both deal with lifting and adjusting loads, there are significant differences between the psychophysical (isoinertial) technique and the Kroemer-McDaniel (isoinertial) protocol in their procedures and the use of the data collected in these tests. For our purposes, we designate the

Jan 1, 1998

During the past 10 years, research work to improve the survey of stope limits after mining supplied useful information, particularly for large open and sub-level stopes, regarding waste dilution caused by differences between final production limits and planned stope limits. However, the efficiency of this information to improve the overall efficiency of the mining process is limited, given that the impacts of inaccurate ore limits leading to dilution (sensu stricto) and ore losses haw not been considered adequately. Geological/sampling information should be optimized in terms of quantity and/or quality for the deposit characteristics and the constraints of the mining method used. As industry practice regarding the definition and recording of dilution is flexible and as ore losses frequently am poorly doc-umented, the use of these terms is reviewed. Simulations were carried out on two typical narrow gold deposits, based on an earlier study at Inco's Thompson mine regarding the impacts of wide to narrow delineation grids using recorded variations of vein geometry. Vertical to sub-vertical deposits of 3 m and 6 m widths mined by long-hole open stopes were considered. Deposits delineated by 7.5 m (25 ft) grids were taken as a base case and the impacts of less accurate limits in conjunction with long hole mining based only on a 15 m (50 ft) delin¬eation grid. The impacts of overestimation (dilution) and underestimation (metal losses) were considered separately, using the average error of +/-0.62 m recorded in the Inco study. The impact of less accurate information was much stronger on the narrower deposit: both dilution and ore losses diminished the net present value of the 3 m wide deposit by 39.27% and 45.41%, respectively, and that of the 6 m wide deposit by 12.84% and 11.55%. As the impacts of ore loss and those of dilution on the net present value am almost identical in these simulations, we could, in practice, assume a 50/50 split between dilution and ore losses in such a case. These results do no take into account the effects of local optimization measures taken at the mine production level (blasthole sampling near stope limits, etc.) The potential impacts of geological and sampling uncertainties in terms of lower NPV justify and require more systematic efforts to ensure that deposit appraisal work and feasibility studies are adequate to optimize company revenue and competitive returns to investors from mine production activities.

Jan 1, 2005

By M. Vallée

"RÉSUMÉDurant les 10 dernières années, les améliorations dans le contrôle des limites des sautages et les arpentages après l’exploitation ont fourni des informations préalables et post facto sur les limites minières réalisées par rapport à la planification, plus particulièrement pour les grands chantiers exploités par les méthodes chambres vides et sous-niveaux. Cependant, peu de recherches ont été faites pour améliorer la délimitation du minerai, l’estimation des ressources et des réserves, les études de faisabilité et la récupération minière dans d’autres types de gisements. Il faudra d’abord expliciter les impacts des lacunes, en quantité et en qualité, de l’information géologique et de l’échantillonnage sur le rendement de l’extraction. Il faudra également remplacer la notion de « dilution générale » par celles de « dilution » et de « perte de minerai ». Les impacts des lacunes observées sur l’exploitation des gisements d’or filoniens sont démontrés par une simulation basée sur des résultats d’études antérieures à la mine Thompson d’Inco, des impacts financiers de densités d’information moindre et par des analyses des valeurs actualisées nettes (VAN) de deux cas types.Deux gisements similaires, de 3 m et 6 m de largeur et exploités par chambres vides sont étudiés, en prenant pour cas de base la délimitation sur maille de 7,5 m (25 pi.) et en étudiant les impacts des limites moins précises supportées par une maille de 15 m (50 pi.), en assumant que ces imprécisions moyennes de +/- 0,62 m se transforment soit en surestimation (dilution), soit en sous-estimation (pertes de minerai).Les impacts d’information moins précise ont été beaucoup plus forts sur le gisement plus étroit : autant la dilution que les pertes de métal ont diminué les VAN du gisement de 3 m de 39,27 % et 45,41 % respectivement, tandis que les VAN du gisement de 6 m ont été diminuées que de 12,84 % et 11,55 %.Ces impacts des valeurs actualisées nettes diminuées justifient et exigent des efforts plus systématiques pour s’assurer que les travaux de mise en valeur et les études de faisabilité soient réalisés de façon requise pour optimiser l’exploitation et offrir aux investisseurs des rendements compétitifs par rapport à ceux des autres secteurs industriels. ABSTRACTDuring the past 10 years, research work to improve the survey of stope limits after mining supplied useful information, particularly for large open and sub-level stopes, regarding waste dilution caused by differences between final production limits and planned stope limits.However, the efficiency of this information to improve the overall efficiency of the mining process is limited, given that the impacts of inaccurate ore limits leading to dilution (sensu stricto) and ore losses have not been considered adequately. Geological/sampling information should be optimized in terms of quantity and/or quality for the deposit characteristics and the constraints of the mining method used. As industry practice regarding the definition and recording of dilution is flexible and as ore losses frequently are poorly documented, the use of these terms is reviewed.Simulations were carried out on two typical narrow gold deposits, based on an earlier study at Inco’s Thompson mine regarding the impacts of wide to narrow delineation grids using recorded variations of vein geometry. Vertical to sub-vertical deposits of 3 m and 6 m widths mined by long-hole open stopes were considered. Deposits delineated by 7.5 m (25 ft) grids were taken as a base case and the impacts of less accurate limits in conjunction with long hole mining based only on a 15 m (50 ft) delineation grid. The impacts of overestimation (dilution) and underestimation (metal losses) were considered separately, using the average error of +/-0.62 m recorded in the Inco study.The impact of less accurate information was much stronger on the narrower deposit: both dilution and ore losses diminished the net present value of the 3 m wide deposit"

Jan 1, 2005

By S Oliver, D Willingham

"Although a resource model is central to the mineral resource value estimation process (Glacken and Snowden, 2001), creating it is a labour intensive task and in the end it is a limited representation of the actual orebody. Creating this resource model is a labour intensive task that requires inputs from geology, mining, metallurgical and commercial disciplines. It requires thousands of samples from hundreds of drill holes to be verified, grouped in geological domain, interpolated and then valued. Even after all this effort, a model is only an estimation; therefore further significant effort has gone into quantifying uncertainty in the model and subsequently risk around any value estimates derived from the model. Further risks to model quality are introduced through poorly conditioned data or incorrect assumptions.This paper proposes the application of machine learning to automate the resource model development. Machine learning is applied to the traditionally manual tasks of geological formation, domain identification and validation of the block model mineralogy. Through automation the resource estimation process can be accelerated, allowing more drill holes or a larger resource body to be processed in a given time frame, or allowing the process to be more agile to changes in input data and assumptions. A case study based on drill hole data from a Western Australian iron ore deposit (Government of Western Australia, Department of Mines and Petroleum, 2015) is used to demonstrate the application of machine learning in this process.CITATION:Oliver, S and Willingham, D, 2016. Maximise Orebody Value through the Automation of Resource Model Development Using Machine Learning, in Proceedings The Third AusIMM International Geometallurgy Conference (GeoMet) 2016, pp 295–302 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 15, 2016

By F Phillips, E Deucker, J Kato

It is challenging to determine the optimal mining strategy to maximise value when a mine plan includes mineralised material which does not meet the full cost of mining but has value exceeding the variable cost of mining (‘marginal material’). Mining too much marginal material dilutes the operation’s head grade and undermines profitability. Conversely, leaving too much metal behind fails to achieve the full value of the endowment. Finding the balance between the two will maximise value delivered from the asset. This dilemma has traditionally been approached by calculating a suite of cut-off grades to define the grade of material that meets economic targets for an orebody (or zone within) and should therefore be included in the mine plan. A problem with this approach is that individual mining blocks within the zone have different costs, value and confidence levels, with each having economic co-dependencies between them. Blocks of identical grade but different mining contexts will have distinctly different economic outcomes. For example: • a block may be a low-grade increment or shell around a higher-grade stope; • a block may be available to be mined when there is spare mining capacity; or • a block may defer access to significantly higher cash flow material. Each of these circumstances would result in different value from the blocks if included as part of the mine plan. Cut-off grades calculated as averages over zones of material, or in isolation from schedule-based information, are missing the entire mining context that determines the ultimate economic outcome of making a mining decision. There are tools and techniques available to address this for open pit mining, however underground mining has different nuances. There are more intricate economic and physical dependencies between potential ore blocks, and importantly, the decision to leave material in situ can be made at any time. The case study in this paper explores a new methodology for underground mining, targeted towards strategic planning applications. It examines whether the economic logic usually applied via cut-off calculations can be implemented in the schedule so that optimal economic decisions can be made on a block-by-block basis, taking into account the mining context of the block, with the objective to maximise asset value. The results of the case study scenarios in Table 1 demonstrate the magnitude of the potential improvement in NPV that can be attained by a block by block valuation approach

Mar 15, 2021

By G. R. Morin, A. R. Leggett

"In 2010, closure of the Kidd metallurgical plants along with structural changes to the historical ratios of payable metals led to a review of the operating targets at the Kidd concentrator. Previous operating data indicated that there was potential for significant revenue improvements by lowering the copper concentrate grade. An extended plant trial was carried out in 2011 to quantify the revenue impact of lowering the copper concentrate grade from 27% to 25% Cu, as well as identifying the lower practical grade limit. Analysis of plant trial results indicated that the process change increased net concentrate revenue by $2 M to $3 M per year at current metal prices. Increased revenue was primarily associated with higher silver recovery to the copper concentrate while the marginally improved copper recovery was essentially offset by increased zinc losses and concentrate treatment and freight charges.Details of the plant test work results are discussed in this paper along with the financial analysis of the revenue improvements. Key areas covered are major revenue components of the concentrate contracts, validation of revenue improvements through plant testing and sensitivity of the improvements to market and plant variables.INTRODUCTIONThe Xstrata Copper Kidd concentrator commenced operations in 1966 treating ore from the nearby Kidd volcanogenic massive sulphide deposit in Timmins, Ontario (Thompson, Amsden, & Chapman, 1978). Over the years, up to four separate grinding-flotation divisions were in operation with throughput rates peaking at 4.5 million tonnes per year of ore in the mid 1980’s (Kelebek & Francis, 1998). As the mine production decreased and grinding throughput increased operations were gradually cut back and since 1998 two grinding-flotation divisions have been in operation treating Kidd mine ore. One of the idled divisions has been intermittently used for custom milling operations, most recently the Ni/Cu Montcalm deposit, which was milled from 2004 to 2009. Current budget throughput rates, limited by mine production, are 2.4 million tonnes per year of ore grading 1.7% Cu, 4.3% Zn and 55 g/mt Ag. Concentrate production rates are 150,000 tonnes per year of copper concentrate and 155,000 tonnes per year of zinc concentrate. The Kidd mine is the deepest base metal mine in the world with mining operations down to 9600 ft.In 2010 the copper and zinc plants at the Kidd metallurgical site, which previously treated the concentrate production from the Kidd concentrator, ceased operation due to a variety of economic and market factors. Following the metallurgical plant closures the Kidd concentrator products were shipped off site for treatment. The majority of the copper concentrate is now processed at the Xstrata Copper Horne smelter in Rouyn-Noranda, Quebec with the majority of the zinc concentrate processed at Canadian Electrolytic Zinc Ltd. in Valleyfield, Quebec."

Jan 1, 2013