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By S. C. Richardson

The electrowinning (EW) circuit involves using DC current in the electro-deposition of metal ions dissolved in an electrolyte solution. This paper reviews a method to generate an electrical model of an EW circuit and the resulting DC current flow and voltage distribution within the circuit. Topics include: calculations of electrolyte conductivity and the equivalent electrical resistance of the electrolyte within the electrolyte pipelines connected to the EW circuit; the breakdown and classification of the EW cell voltage into thermodynamic and ohmic potentials; methods to calculate current flow and voltage distribution throughout the EW circuit; and the development of an interactive visual computer model to simulate an actual EW circuit. Stray current can pose an electrical shock hazard to employees in electrolytic tankhouses. Increased awareness and understanding of the voltage distribution and potential stray current in an EW circuit can help mitigate the risk of tankhouse workers becoming exposed to electrical shock hazards. INTRODUCTION Copper electrowinning tankhouses include rectifiers that deliver large amount of DC current to a series of electrolytic cells consisting of electrolyte solution and electrode pairs. The design also includes metallic conductors in the circuit which include the rectifier’s positive and negative busbar, inter-cell busbar, and the anode and cathode frames. The electrolyte acts as an electrical conductor that transfers current via ions between the anodic and cathodic electrodes within the cell. Electrolyte, as a conductor, is also capable of conducting DC current outside the main EW circuit through the connected electrolyte feed and return pipe systems. DC current that follows a path other than the intended path through the series of electrolytic cells is referred to as “stray current”. Developing an electrical model enhances understanding of voltage distribution and current flow in an EW circuit. This paper discusses the derivation of the various representative components of the model as well as the output from the model. An interactive visual computer simulation incorporates the model output in order to aid in the understanding of voltage distribution and current flow throughout the EW circuit.

Jan 1, 2019

By F. A. Auld

"During the period 1958-1980 seven deep shafts were sunk from the surface in the UK. From 1977 until 1989 sixteen deep shafts were sunk in the UK, all concrete lined. Subsequently from 1989 up to the present day no deep shafts have been sunk in the UK, a period of 30 years. Developments in the design and construction of deep shaft linings in the UK therefore were predominantly carried out throughout the period 1977–1989. This knowledge is recorded in several published papers by the author over that period which will be referred to in the paper. INTRODUCTION The paper includes a review of past precedent for the design of deep shaft concrete linings in the UK and sets out the principles for future designs, including the design load to be adopted and Factors of Safety (SFs) to be used in the design. Hydrostatic pressure increases with depth from the surface and therefore the concrete lining must increase in thickness and strength throughout its depth to resist the pressure. High strength concrete mixes were developed to enable economic linings to reach greater depths and these will be detailed. The ground conditions to sink through in the UK consist of predominantly competent (self-standing when excavated) rock, which can be water bearing, and hence a hydrostatic pressure resisting lining is required. The paper describes the construction measures needed to produce a dry shaft by way of pour length, joint detail, backwall grouting and joint sealing details. PAST PRECEDENT FOR UK DEEP SHAFT SINKING AND LINING Seven deep shafts were sunk from the surface in the UK from 1957–1977 over a 20-year period (Auld, 1987). These are listed in Table 1. Three shafts were for British Coal, two were for Cleveland Potash Ltd and two were for the Central Electricity Generating Board. The British Coal shafts were 5.486 m and 7.315 m internal diameter to depths of 550 m and 780 m, respectively. At 5.5 m internal diameter and a depth of 1150 m, the two Boulby Mine shafts for Cleveland Potash Ltd are currently the deepest shafts in the UK. The two Central Electricity Generating Board shafts of 5 m and 10 m internal diameter, to depths of 255 m and 440 m, respectively, were ventilation shafts for a pumped storage scheme. Six of the shafts were sunk by Thyssen (GB) Ltd and one by Cementation Mining Ltd."

Jan 1, 2019

By M. Z. Emad, M. U. Khan, A. ur Rehman

Stemming plugs are used as an accessory in drilling and blasting industry. They supplement conventional stemming material (i.e. drill cutting) with retention of gases in the drill hole that result in better fragmentation of rock and improved explosive utilization. Better fragmentation and explosive utilization along with economic benefits enhance efficacy of the project. Incorporation of stemming plugs in drilling and blasting industry is in exercise for about thirty years. The efficiency of stemming plugs is a key factor in their rising induction in mining industry. Different studies have been conducted on the economic viability of these plugs. The focus of this paper is the selection of stemming plugs for a bench blasting operation and the examination of blasting operation efficiency based on ergonomics and human factors. Three types of stemming plugs were evaluated during multiple blasts at a limestone quarry. Their performance was evaluated using split desktop software. Based on the human factors plastic molded stemming plug (stemming plug-1) was found to be the best among all selected plugs. INTRODUCTION Surface mines and quarrying operations heavily depend on bench blasting to achieve high volumetric targets and production demands. Use of explosives is the most economic method adopted for rock breakage and movement. Extensive research had been performed to enhance explosives energy utilization. This include changing the drilling pattern, application of air decks, stemming plugs, and use of other supportive devices. Optimization of explosive energy has been a major concern in those mines that rely on drilling and blasting operation to carry out main production. Due to continually increasing cost, efficient mining demands, and associated hazards of using explosives, it is necessary to utilize explosives at their best. Efficiency of a blasting operation is generally evaluated by the fragmentation size. From post blast conditions at blasting site to efficiency of processing plant, anything can be affected due to improper blast resulting in bad fragmentation. Better fragmentation is necessity for better operation of the system along with consideration of the safety of operation, environmental issues and related costs. For developing a blast design, an engineer should always attempt to minimize the cost of operation without compromising with acceptable technical and safety necessities [1]. Blast performance is dependent on many different factors associated with the process. These factors have been gauged as blast geometry, explosive specification, and rock mass properties [2]. Similarly, in 2013 Mohammad et al. concluded that fragmentation dynamics depends on blast hole diameter, primer, spacing, stemming, burden, and delay timing for blasting operation [1]. At the work site where this study was performed, blast holes were filled by drill cuttings and muds to help in confining gases. In a blast hole, gas confinement helps in better rock fragmentation along with reduced flyrocks. This practice is most common in many countries including Pakistan. Due to lack of homogenized particle size, the effectiveness of confinement is variable and majorly based on the size of stemming. Extremely small sized drill cuttings often result in hole blow out along with flyrocks from the collar. This improper blast energy will result in poor blast, bad impacts on environment, and reduced operational efficiency. In addition to the stemming material, the length of stemming is also a prime consideration. The decrease in stemming length will increase the quantity of explosive in the blast hole and an increase in stemming material generally improves stemming performance.

Jan 1, 2019

By N. Snyders, I. Mwandawande, S. M. Bradshaw, M. Karimi

"Computational fluid dynamics (CFD) was applied to predict the average gas holdup and the axial gas holdup variation in a 13.5 m high cylindrical column 0.91 m diameter. The column was operating in batch mode. A Eulerian-Eulerian multiphase approach with appropriate interphase momentum exchange terms was applied to simulate the gas-liquid flow inside the column. Turbulence in the continuous phase was modelled using the k-_ realizable turbulence model. The predicted average gas holdup values were in good agreement with experimental data. The axial gas holdup prediction was generally good for the middle and top parts of the column, but was over-predicted for the bottom part of the column. Bubble velocity profiles were observed in which the axial velocity of the air bubbles decreased with height in the column. This may be related to the upward increase in gas holdup in the column. Simulations were also conducted to compare the gas holdup predicted with the universal, the Schiller-Naumann, and the Morsi-Alexander drag models. The gas holdup predictions for the three drag models were not significantly different. IntroductionColumn flotation is an important concentration technology that is used in the mineral processing and coal beneficiation industries. The growing interest in the use of column flotation in mineral processing has been attributed to the simpler flotation circuits and improved metallurgical performance compared to conventional flotation cells (Finch and Dobby, 1990). Flotation columns have also found other applications outside mineral processing, such as de-inking of recycled paper (Finch and Hardie, 1999).In column flotation, a rising swarm of air bubbles generated by means of air spargers is employed to collect the valuable mineral particles and separate them from the gangue minerals in a countercurrent process. Wash water, which is continuously fed at the top of the column, is used to eliminate entrained particles and stabilize the froth. The column volume can be divided into two sections – the collection zone in which the bubbles collect the floatable mineral particles, and the cleaning zone (or froth zone) where product upgrading is enhanced through the removal of unwanted particles entrained in the water mixed with bubbles from the collection zone."

Jan 1, 2019

By E. Talebi, P. Roghanchi

"This study aims to discuss personal and behavioral parameters contributing to heat-related accidents. To study the prevention approaches, it is essential to look at the various factors at the individual, management, and industrial levels. Qualitative risk analysis methods are used based on the public data in the USA to find the impact of the environmental and behavioral parameters on the heat stress and heat strain of mine workers. The effects of individual factors such as worker experience, age, fitness, lifestyle, acclimation, and more are illustrated. The significance of factors such as the roles of the mine management and regulatory agencies are addressed.INTRODUCTIONOccupational hazards in mining industry endanger the health and safety of the workers and productivity loss for mining companies. Among those hazards is the exposure to excessive heat that has adverse effects on the health of mine workers. The intensity of heat load is a significant factor in fatal and non-fatal heat-related accidents. Severe environmental conditions combined with a high level of activity in the mining industry increases the vulnerability of individuals to health and safety hazards (Ngwenya, Oosthuizen, Cross, Frimpong, & Chaibva, 2018). Lack of regulations and legislation addressing the heat issue in the mining industry (i.e., underground and open pit mines, processing plants) also raises widespread concerns.Along the climate change, various sources contribute to the overall heat load at a mine site (Hanna, Kjellstrom, Bennett, & Dear, 2011; Kjellstrom, Lemke, Hyatt, & Otto, 2014). In underground mines, the majority of heat comes from different sources such as geothermal gradient, equipment, auto compression, groundwater, and blasting (Maurya, Karena, Vardhan, Aruna, & Raj, 2015). On the surface, however, the majority of heat load can be sourced to solar energy. The exposure to excessive heat might result in heat-related illnesses, fatigue and fainting, loss of concentration, performance reduction, and production loss (Kocsis & Sunkpal, 2017; Roghanchi, 2017). The severity of the effects depends on duration and level of exposure and the response of an individual to heat stress. The response of an individual is influenced by different factors including age, gender, fitness, lifestyle, experiences, and past medical history."

Jan 1, 2019

By S. Frimpong, E. Gbadam, W. Nyaaba, G. Somua-Gyimah

Recent studies towards dragline excavation efficiency have focused on incrementally achieving automation of the entire excavation cycle. Initial efforts resulted in the development of an automated dragline swing system, which optimizes the swing phase time. However, the system still requires human operation for collision avoidance. For full dragline autonomy, a machine vision system is needed for collision prevention and big rock handling during the ‘swinging’ and ‘digging’ phases of the excavation operation. Previous attempts in this area focused on collision avoidance vision models which estimated the location of the bucket in space in real-time. However, these previous models use image segmentation methods that are neither scalable nor multi-purpose. In this study, a scalable and multi-purpose vision model has been developed for draglines using Convolutional Neural Networks. This vision system averages 82.6% classification accuracy and 91% detection in collision avoidance. It also achieves an 87.32% detection rate in bucket pose estimation tasks. In addition, it averages 80.9% precision and 91.3% recall performance across terrain recognition and oversized rock detection tasks. With minimal modification, the proposed vision system can be adjusted for other automated excavators. Keywords: Surface mining, dragline bucket, deep learning, machine learning, oversized rock, convolutional neural network, machine vision, object detection, earthmoving equipment. INTRODUCTION Complete excavator automation is widely regarded as the next phase in efforts toward improving earthmoving efficiency. Earthmoving often involves complex, forceful interactions between an excavator and the ground. The nature of these interactions depends largely on the type, properties and physical characteristics of the earth material. This process is further complicated by the random occurrence of tree roots, boulders and other such obstructions. Therefore, the ability of a controller to detect changes in the operating conditions, adjust the digging strategy and respond in real-time is of utmost importance. Early studies into autonomous excavation identified some key performance criteria which included the following [1]: • The autonomous excavator must be able to work in any type of earth material. • Its excavation accuracy must be within 50mm. • It should be able to handle different surface and underground obstacles autonomously. • It should be able to operate at the speed of the average operator in any condition. • Its operation should be capable of safe integration with other site systems.

Jan 1, 2019

By O. Mejia

"As study was undertaken to investigate the relationship between mill liner wear and grinding energy efficiency and to assess the opportunity to control mill speed to improve energy performance as mill liner wears. The study involved analysis of plant operating data and liner wear profiles over time. The operating work index was found to be an effective tool to identify the decrease of grinding efficiency along the liner service. The plant data show that the mill grinding efficiency as determined by monitoring the operating work index, and throughput decreases as liner wears. However, efficiency and throughput can be maintained by either changing the liner more frequently or by increasing the mill speed. Two scenarios for increasing mill throughput were analyzed. The first one involves changing the liner before the efficiency drops and the second one requires increasing the mill speed from 75% to 79%. Both scenarios show financial benefits that can be used in mill operations.INTRODUCTION The analysis of the mill power draw is used in this study as a key parameter for comparing mill liner profile and its effect on mill performance. Current technology and simulation software used in the mining industry do not factor in the impact of liner wear profiles in grinding efficiency and mill throughput (Toor, 2013). To address this gap, this study provides an analysis of key factors affecting mill throughput and grinding efficiency: mill speed, ball loading, and liner profiles. Several years of plant mill operating data, published research papers, and MillTraj® simulation results are used to demonstrate that face angles impact grinding efficiency and throughput and that mill speed can be used to overcome this problem. Moreover, the liner profile wear analyses clarify a mismatch between actual and estimated power draw calculations."

Jan 1, 2019

By L. Garycki, P. Poks, R. Urbaniak, T. Gabka

In 2016 at Głogów Copper Smelter KGHM Polska Miedź S.A., after shutting down the shaft furnaces new FSF was launched. Currently, it is the only smelter with two direct to blister flash smelting furnaces in operation. The main difference between the vessels are the dimensions and production capacity. The new furnace design output is about 30% higher compared to the old flash smelting furnace working for 40 years. The paper presents a comparison of both FSFs operation, differences in technology and the way to optimize the process to maximize efficiency. INTRODUCTION KGHM Polska Miedź S.A. is the only company in Poland (eastern part of Europe) that extracts and processes copper ores. It is located in the south west of Poland and this region is called the "Copper Basin”. KGHM PM S.A. consists of three underground mines: Lubin, Polkowice-Sieroszowice, Rudna, concentrator plants, a tailings pond plant, three copper plants: Legnica, Głogów and Cedynia wire rod plant in Orsk. The publication relates to the operation of the furnaces in Głogów (the plant in Głogów consists of Głogów 1 and Głogów 2 which form integrated smelting and refining facilities). In 2016, after closure of the shaft furnaces technologies in the Głogów 1, a new FSF commenced operation. The simplified process flow sheets are given in Figure 1 and 2. Technological schemes are similar, but due to the differences in them, they will be discussed separately. Głogów 1 Enriched ores are discharged and directed to the feed preparation plant. Delivered concentrates are built up on the three bedding aisles in appropriate proportions. The concentrates are distributed in even layers in one of the three aisles. Aisle emptying is done with a blending machine which cuts across the concentrate layers and collects thus blended material, sending it via belt conveyors to a storage bin. From the cross- blending machine and conveyor assembly, the mixture of concentrates is directed to drying in the steam dryer. Part of the charge is directed to the FSF and part to the roaster for carbon elimination. Known in the world roasters are used to eliminate sulfur compounds or arsenic. In the case of KGHM deposits, the aim of the process is to reduce the energy value of the feed by eliminating carbon and converting sulfur from sulfide to sulfate. Changing the nature of exothermic to endothermic material is crucial in the case of furnace operation with design performance. Roaster is a separate furnace with an innovative character.

Jan 1, 2019

By A. Karzhau, B. Aldamzharov, Y. Tussupbekov

"This case study produced for the metal mine operating in Kazakhstan, which has ventilation controlled manually and working continuously at maximum speed to keep air flowing through each part of the mine 24 hours a day. The mine auxiliary ventilation has been designed with the maximum length of ducting systems with extra air volumes to compensate for potential leakage. The mine provided auxiliary ventilation to all production stopes even to those, which for various operating reasons become temporarily inactive. However, mine safety has to consider risks associated with the deficit of fresh air for mine gases dilution and removal. The mine needs additional investments for re-designing ventilation network and therefore needs to estimate financial risks associated with it. The aim of this study is to conduct risk analysis for ventilation network optimisation processes at the operational metal mine in Kazakhstan. Mine ventilation surveys were undertaken for validation of the computerised ventilation model. The mine ventilation network modelling has been undertaken to be compliant with Kazakhstan mining regulations. Results of the mine ventilation optimisation processes were analysed in terms of financial risks for investments into reconstruction of the ventilation shafts, extension of the ventilation network, variable speed fans and automated regulators. Savings were resulted from improving ventilation system efficiency by reduce airflow in areas without mining activity and redistributing existing ventilation capacity. Based on savings derived from the optimised ventilation network an additional NPV has been received. Stochastic modelling use Monte Carlo simulation producing representation of possible outcomes for the additional NPV considering electricity pricing as a major variability. As result probability distribution for the optimised ventilation network demonstrates the additional NPV in a range of 1.6 - 2.2 million US$ in the corresponding defined value of 90%. Moreover, there is only 5% chance that the additional NPV to be less than 1.5 million US$. Conducted risk analysis for ventilation network optimisation processes gives considerations on volume of investments into the mine ventilation system.INTRODUCTIONCosts for mine ventilation are typically increasing when mines expand to deep levels. Some underground mines in Kazakhstan still use continuously operating fans at its maximum capacity. This is not helping to save energy and costs that are in many cases very significant. This case study produced for the metal mine operating in Kazakhstan, which has ventilation controlled manually and working continuously at maximum speed to keep air flowing through each part of a mine 24 hours a day. The mine auxiliary ventilation has been designed with maximum length of ducting systems and extra air volumes to compensate for potential leakage. The mine provided auxiliary ventilation to all production stopes even to those, which for various operating reasons become temporarily inactive. However, mine safety has to consider risks associated with the deficit of fresh air for mine gases dilution and removal from the mine workings (Sabanov, 2019). The mine needs additional investments for re-designing ventilation networks and therefore need to estimate financial risks associated with it. The aim of this study is to conduct risk analysis for ventilation networks optimisation processes at the operational mine in Kazakhstan. Mine ventilation surveys for validation of the computerised ventilation model have been undertaken"

Jan 1, 2019

By A. Hutwalker, T. Plett

"Blast fumes are one of the major sources for NOx in underground mines. In 2016, the Occupational Exposure Limits for NOx in Germany were reduced by 90%. The task is to keep re-entry times after blasting at the present level and thereby ensure an economic operation of the mine. Measurements show that using ventilation solely is not sufficient. A new approach to reduce NOx from blast fumes is described. Laboratory experiments were performed and a reduction of NOx was observed. Measurements were conducted at an explosive factory’s testing facility and in an underground mine at blasting operations. INTRODUCTION Whether for production of ore or for drift development, the processes in an underground mine are always the same: 1) fracturing the rock, 2) loading the pile, 3) support the resulting cavity. The fracturing of the rock can be done mechanically, by cutting, for example by using continuous miners or tunnel boring machines, or by drilling and blasting. Drilling and blasting is more flexible than mechanical driving, as it can be used independently of the rock strength. Furthermore, this driving method allows a more flexible reaction to changes in geological and/or operational conditions. Thus, underground mining and blasting are almost inextricably linked with each other, and regardless of the size and level of technology of the mine, blasting is hardly avoidable. Depending on the mining method chosen and the existing framework conditions, such as the mechanical properties of the rock or the availability of machinery, blasting is an integral part of daily operations. Even in mines where blasting is not part of regular operations, it cannot be completely avoided. The blast fumes generated by the explosions are, in addition to diesel engine emissions, one of the two main sources of emissions of nitrous gases underground. Up to now, state of the art for dealing with blast fumes is ventilation. The fumes are diluted with fresh air and ventilated out of the mine. This treatment of blast fumes is more than sufficient with the former German Occupational Exposure Limits (OEL). The adaption of the OEL in Germany to the recommendations of the Committee on Occupational Exposure Limits (SCOEL), a committee of the European Commission, leads to big challenges for the mining industry in reducing the concentrations of these gases at the workplaces underground. On its meeting on May 2nd and 3rd 2016, the Committee on Hazardous Substances (Ausschuss für Gefahrstoffe, AGS), an advisory body of the German Federal Ministry of Labour and Social Affairs (BMAS), advised the BMAS to copy the OEL, as recommended by the SCOEL, to the Technical Rules for Hazardous Substances (TRGS 900). BMAS followed this advice, consequently the new OEL where published on November 4th 2016 and became part of the TRGS 900. Table 1 shows the former and new OEL for Germany. These new OEL became law after publishing, but two exceptions were made: For tunneling operations, they became law on October 31st 2017, for mining operations they will become law on October 31st 2021 (BAuA, 2004, 2016, 2017)."

Jan 1, 2019

By M. Deutsch

A common casualty of modern open pit mine optimization is the assurance that the resulting design is actually achievable. Optimized mine plans that consider value and a bare minimum of precedence constraints do not, in general, translate into practical, operational mine designs that can really be used in the field. Ultimate pits may come to a sharp point at the bottom. Schedules may require taking small parcels of material from many disparate areas of the pit in a single period. And grade control polygons may be ragged, narrow, and not minable with realistic equipment. In this paper all of these problems are addressed by encoding these three fundamental open pit mine optimization problems as maximum satisfiability problems. Maximum satisfiability provides a useful framework for problems that are non-linear and may guarantee the optimality that metaheuristics cannot. INTRODUCTION Open pit mines are large and complicated operations, which require significant initial and ongoing investment. A single mine may employ hundreds or thousands of people and creates substantial benefits for both the surrounding and global communities. Planning an operation of this magnitude requires a lot of effort and cannot be done optimally by hand. Mining engineers today rely on many techniques from operations research to guide decision making, and to maximize the value of the mine. Techniques including linear programming, metaheuristics, and network algorithms are all used to solve different problems throughout mine design and optimization. However, techniques based on Boolean satisfiability, including the optimization extension of maximum satisfiability, have not seen much use in mining despite being used in many other fields including electronics design, scheduling, and artificial intelligence. We claim that maximum satisfiability is a useful and meaningful way of expressing and solving optimization problems in mine planning. However, this technique does not come without challenges; the problem is NP-hard, and modern techniques for solving these problems can struggle with the number of variables and clauses that a typical mining problem requires. Also, some of the constraints that mining must contend with do not fit nicely into the satisfiability paradigm. Despite these challenges, we believe that maximum satisfiability provides a useful framework for considering operational constraints and expressing some problems that other techniques can't easily express. Maximum satisfiability can be solved exactly, yielding the true optimal answer which metaheuristics cannot guarantee. In this paper we show the applicability of maximum satisfiability to mining by specifying three fundamental problems in open pit mine optimization. We show how these fundamental problems can be formulated and how they can be extended within the framework of maximum satisfiability. For the remainder of the paper we first give a brief background on common elements between the three problems and an introduction to satisfiability. Then we tackle the three problems; the Ultimate Pit Problem, the Block Scheduling Problem, and the Grade Control Polygon Problem. In the Ultimate Pit Problem we specifically show how to extend the problem to support a minimum mining width – a coveted result for many mining engineers. We then discuss some of the complexities and shortcomings of this framework and possible research directions.

Jan 1, 2019

By Zhuo Chen, Jun Zhou, Jiemin Zhou

Samples of the reacting suspension particles were taken from the reaction shaft of a commercial flash furnace at different heights. The processes of matte and slag formation and morphology of the samples in the shaft were observed and analyzed. Variations of matte and slag composition in the shaft were investigated. By analyzing the samples taken from the settler at different locations, settler processes were studied such as slag formation, agglomeration and separation of matte-slag phases, Fe3O4 reduction, copper-in-slag reduction etc. In accordance with the characteristic of matte and slag formation the flash furnace can be divided into 6 functional regions, and a mechanism model called as “Micro bath process” is proposed for matte and slag formation in a commercial flash furnace. INTRODUCTION During flash smelting, concentrate particles scattered in the reaction shaft complete their oxidation reactions in suspension. The droplets produced fall into the settler beneath with a small amount of un-melted solid particles, agglomerate and separate into matte and slag layers which are tapped out from their tap holes separately. With respect to the extent of slagging reaction in the reaction shaft, there isn’t any commonly accepted understanding in literature (Kimura, Ojima, Mori & Ishii, 1986; Kemori & Kondo et al.,1988; Kemori, Denholm & Kurokawa, 1989; Fagerlund &Jalkanen, 2000). Some assume that the slagging reaction in the shaft can be ignored and some propose that all slag is produced in the shaft. With regards to the characteristics of the physical-chemical changes of concentrate particle in the reaction shaft, although there are some research results reported from pilot scale and commercial scale experiments (Kemori, Ojima & Kondo, 1988; Hagni, Vierrether & Sohn, 1988; Jorgensen, Taylor, Dickinson & Stevenson, 1992; Swinbourne, Simak & Yazawa, 2002; Kivinen & Tuominen, 2005), they are generally limited themselves to theoretical analysis and calculations. No studies are published concerned with the processes of matte and slag generation in the shaft, quantitative analysis of the processes, variations of matte and slag composition and morphologies of the melts along the axis of the shaft, etc.

Jan 1, 2019

By J. Sammut, T. G. Joseph, R. Andreev, M. Curley

"Decisions regarding mining method and subsequent choice of excavator must take into account the quantitative factors of productivity, availability, and capital, operating, and maintenance costs. Dig performance was compared between a hydraulic excavator (HE) and electric rope shovel (RS) of similar size class (i.e., use the same amount of energy to excavate a unit quantity of the same material from a mining face), focusing on the quantitative factors above. Machine performance, normalized to bucket capacity, was assessed as a function of hours of service (60,000 hours per machine). Results showed higher production rates for the HE of similar capacity and age. In general, the RS exhibited lower service cost per volume capacity, though it had higher capital cost. The RS became more cost effective than the HE of similar capacity within approximately 30,000 hours of service. ¦ KEYWORDS Cost, Electric rope shovel (RS), Excavation, Hydraulic excavator (HE), Surface mining RÉSUMÉ Les décisions concernant la méthode minière à adopter et le choix ultérieur d’un engin d’ex¬cavation doivent tenir compte des facteurs quantitatifs de la productivité, de la disponibilité ainsi que des coûts d’investissement, d’exploitation et d’entretien. Cette étude compare les performances en termes d’excavation d’une excavatrice hydraulique et d’une pelle électrique à câbles de même taille (qui utili¬sent par exemple la même quantité d’énergie pour creuser une quantité unitaire du même matériau dans un front de taille), en se concentrant sur les facteurs quantitatifs susmentionnés. La performance de la machine, normalisée à la contenance du godet de volume unitaire, est évaluée en fonction des heures de service (60 000 heures d’activité par machine). Les résultats montrent des taux de production plus élevés pour l’excavatrice hydraulique d’une capacité et d’une ancienneté similaires. De manière générale, la pelle électrique à câbles affiche un coût d’entretien par capacité volumique inférieur, mais son coût d’in¬vestissement est plus élevé. La pelle électrique à câbles devient plus rentable que l’excavatrice hydrau¬lique d’une même capacité au bout d’environ 30 000 heures de fonctionnement. ¦ MOTS CLÉS coût, excavation, excavatrice hydraulique, exploitation à ciel ouvert, pelle électrique à cablesINTRODUCTION The most common excavating tools in surface mining operations are the hydraulic excavator (HE) and electric rope shovel (RS). In general, electric RSs are more reliable and cheaper to maintain (P&H Mining Equipment, 2007), but where complicated geological conditions exist, HEs are preferred because they are generally considered to be more mobile."

Jan 1, 2019

By A. J. Rigby

The wear mechanisms of the tuyere lines of Peirce-Smith and Hoboken converters are a subject of endless debate and speculation. However, there has been an indication that one of the primary factors in the deterioration of tuyere lines is due to the batch operation of the converting process. The cooling of the brick in the tuyere line during out-of-stack times leads to shrinkage of the brick and resultant opening of the joints. Subsequently, on re-entry of the tuyere line into the bath, the joints become penetrated with matte and re- expansion of the brick is prevented. It is thought that this regular cycle of exiting the batch for slag skimming and replenishment of matte and re-entering for further blowing leads to a gradual “ratcheting” of the tuyere brick lining. If no expansion allowance is furnished between the endplate and the endwall of the lining, then the generation of damaging compressive stress is unavoidable. INTRODUCTION It is generally accepted industry-wide that the integrity of a refractory lined vessel is determined by an inextricable combination of refractory brick quality, the caliber of the lining design and installation, and the severity of the process inflicted upon the high wear areas of the vessel lining. Rigby has summarized the concentrated attempt, over the past several years, to optimize the microstructural, chemical, and fracture properties of the refractories used in non-ferrous pyrometallurgical vessels. Furthermore, lining design concepts involving larger and longer brick and the extent and positioning of expansion allowance in the joints have been extensively trialed. All these efforts have provided some, but not startling, improvements. The simple fact is that a vessel lining composed of bricks with an optimized microstructure and chemical inertness must installed with a well-conceived design to decompress the lining stresses that are generated by the batch operation of copper converters. REFRACTORY QUALITY Converting is conducted by injecting air into the molten matte and white metal for the very purpose of generating iron and copper oxide slags. These slags must be contained by a refractory which, by its mineralogical formulation, is optimized for chemical inertness.

Jan 1, 2019

By S. Konetschnik, Lux Degel, R. Degel, T. Haslinger

UrbanGold and its partners from Mettop and SMS group have developed a novel technology for the recovery of valuables from all non-ferrous metals’ fractions of WEEE (Waste Electrical and Electronic Equipment), with a special focus on low qualities which are commonly landfilled or incinerated. The technology is marketed under the brand name “UrbanGold Flex” to illustrate the high flexibility in respect of plant design and raw material portfolio, allowing also the processing of low grade material and materials with high caloric values. The process of the technology’s centrepiece, a modified and enhanced bath smelting furnace called “HENRI”, was investigated in numerous pyrometallurgical trials as basis for a scale-up towards a capacity of up to 100,000 tons of WEEE-concentrates per year. All relevant facility components have been evaluated in respect of references, fall-back options, as well as exit strategies (alternatives) to minimize risks when establishing a novel technology. A spin-off within the development is the “UrbanGold Compact” technology which can be seen as optimization of state-of-the-art smelters by using an optimized TBRC as primary processing unit. As an UrbanGold Compact plant will be the first industrial scale reference of UrbanGold’s technologies, it is also explained in the publication. As precious metals recovery is of major (economical) importance when recycling WEEE, major attention has been given to minimize the losses of these valuables. These measures, as well as a sophisticated way of utilizing waste heat as far as possible, result in interesting business cases with gross profits of up to 20% or even more. INTRODUCTION Over the past years, the demand of electrical and electronic equipment has increased dramatically with technological progress. Innovations on the technology of electronic devices led to a shorter usage and lifespan and thus boosted the generation of Waste Electrical and Electronic Equipment (WEEE). WEEE is one of the fastest growing waste streams, growing at 3 to 4% per year worldwide and is expected to reach more than 52 million tons per year by 2021. Currently, about 45 Mt of these wastes are estimated to arise worldwide, and the trend is rising (Baldé et al., 2017).

Jan 1, 2019

By L. Yuan, J. H. Rowland, R. A. Thomas, D. Bahrami

Blasting is the main method of production in many non-coal underground mining operations and produces multiple toxic gases as a result. The Mine Safety and Health Administration (MSHA) requires mine operators to measure the level of toxic gases in mines as frequently as necessary to ensure they are below regulatory safety limits. The current practice uses portable gas monitors to check the concentrations of toxic gases after a fixed post-blast time. This paper studies the application of a gas monitoring system in the return entry of a limestone mine to determine a safe re-entry time. The National Institute for Occupational Safety and Health (NIOSH) conducted such a monitoring program in a limestone mine from September 2016 through May 2018. NIOSH/PMRD (Pittsburgh Mining Research Division) is endeavoring to develop workplace solutions to improve detection of and reduce the risk of hazardous conditions. This study showed that the use of gas monitoring in the return air can be a useful tool at the mine operator’s disposal to detect and reduce the risk of hazardous conditions and also to reliably estimate the re-entry time. INTRODUCTION Underground miners are at risk of overexposure to toxic gases from various sources. Production blasting operations using explosives in underground mines are a source of multiple toxic gases that could be harmful and even fatal to humans if prolonged exposure occurs beyond the safe limits. According to Code of Federal Regulations, Title 30, Part 57.5002 [1], metal and nonmetal mine operators must conduct gas, mist, and fume measurements as frequently as necessary to determine the adequacy of control measures to protect miners from overexposure to harmful gases. MSHA will inspect mine operators’ activities to verify the evidence of such measurements [2,3]. Continuous gas monitoring in the return air is an effective way to make sure that toxic gas concentrations at the working areas are kept below the regulatory limits before the return of miners. Toxic gases can enter the mine atmosphere from different operations, such as welding, combustion, and blasting. The main gases produced from blasting are carbon monoxide (CO), carbon dioxide (CO2), nitric oxide (NO), and nitrogen dioxide (NO2). Other gases are produced in small amounts and, therefore, were not considered for this study. Related regulatory and recommended exposure limits of CO, NO, and NO2 gases are summarized in Table 1 [4].

Jan 1, 2019

By Barry Hope, Javier Rico

"Goldcorp’s Hoyle Pond mine has recently completed several projects which have impacted the overall ventilation system. The project includes extending the Fresh Air Raise (FAR) to depth, development of a new Return Air Raise (RAR), and the installation of new sensors and ventilation controls throughout the mine. Two years after the completion of #2 Winze, the mine is ramping up its ore production to 1,300 tpd (up from 1,100). To ensure safe and profitable production, a reliance upon technology has been needed to provide the required ventilation to the work areas according to the mine plan. To systematically assess the mine’s ventilation in real-time, a significant increase in remote monitoring was required. This included installing additional flow, pressure, temperature and gas sensors. The first step in upgrading the existing monitoring was to purchase airflow and differential pressure monitors and install them in key areas of the mine. This paper presents the process of expanding the monitoring capability at Hoyle Pond mine, the selected equipment and the results of the implementation. The intended goal was to improve safety while minimizing capital and operation costs. The monitors were installed in strategic locations to provide immediate feedback on the impact of changes to the primary ventilation airflows. Comparing the available sensor technologies, installing them and systematically monitoring the ventilation served as a first step towards developing a mine wide Ventilation Control System. In addition, the initial state of the monitoring and control system, with final results and the chosen solution for the mine site, are presented. INTRODUCTION The Hoyle Pond mine is located in the City of Timmins, within the Porcupine mining district, Ontario, which has produced more than 70 million ounces of gold since 1910. Currently, approximately 60% of Porcupine Gold Mines gold production comes from the Hoyle Pond underground mine. Mechanized cut and fill and long-hole mining methods are used to extract the ore. The remainder of the ore comes from the Hollinger Open Pit mine. The Dome Underground mine ceased operations in 2017. Hoyle Pond is currently mining at 1820 meters below the surface and plans to go deeper. The average dry-bulb as wet-bulb temperature underground varies according to the depth and surface conditions which has extremes of –40°C in winter to +30°C in summer. The total airflow volume pushed down from surface is 240 m3/s to accommodate the diesel-powered equipment needed for normal operations. In order to improve the quality of air underground a program has been implemented to use diesel particulate filters, starting with the truck fleet and 6-yard scoops, followed by the intention to implement the rest of the functional diesel-powered equipment and eventually move to electric equipment."

Jan 1, 2019

By G. R. C. Pacheco, D. M. Fonseca, R. Neuböck, G. E. Gonçalves, W. Moulin-Silva

Degradation by sulfur gases is an important lining wear mechanism in the copper industry. SO2/SO3 gases infiltrate into the pores of the basic refractory bricks, combine with their raw materials through expansive reactions and significantly modify the bricks’ structure. This study presents a comparison of the corrosion resistance against sulfur attack of different magnesia-based refractories using a simple, low-cost and quantitative method. Recipes containing sintered MgO grains showed lower resistance to corrosion by sulfur oxides compared to basic grades containing fused grains. Chromium corundum grades were used as a reference and showed very good resistance to this corrosion mechanism under the test conditions. The novel test provided relevant results that agree with what is commonly observed during operation and in previous post-mortem studies. This tool could be very useful in selecting refractory grades for the copper industry and predicting their performance. INTRODUCTION Oxidation of matte sulfides during copper production leads to the formation of sulfur oxides, which infiltrate into the pores of the refractory lining and react with the basic oxides below 1000°C to form magnesium and calcium sulfates. This reaction is accompanied by an expressive volume increase, consequent densification of the refractory microstructure and weakening of the brick bond (Malfliet et al., 2013). Thus, corrosion by SO2/SO3 is an important wear mechanism in the non-ferrous industry and the resistance to this compound should be considered when defining refractories for this application. Previous studies on this matter include expensive and complex testing methods involving pressurized SO2 gas cylinders in order to investigate the corrosion resistance of MgO-containing materials by sulfur oxides (Gerle & Podwórny et al., 2016; Fotoyi & Eric, 2011; Fotoyi, 2009; Podwórny & Piotrowski et al., 2008). The current study was based on a test developed by Pacheco & Fonseca (2018) to assess alkali infiltration and aims to provide an evaluation method of the corrosion resistance of refractories by SO2/SO3 in a simple and quantitative manner. This novel test is useful to compare the corrosion resistance of different refractory recipes and properly select the grades that should be employed in the non-ferrous industry. A case study is provided to illustrate the damage caused by sulfur oxides in refractory linings.

Jan 1, 2019

By Chris Dougherty

"This presentation will provide a demonstration of the methodology used in designing a modern shaft and hoisting plant utilizing a review of the challenges faced and solutions delivered for the Elk Creek Superalloy Materials Project (Niobium/Scandium). The development of the design required the review of complex hydrogeological conditions, mine production requirements, demanding capital expenditure (CAPEX) limitations and tight schedules. Through an extensive review of various alternatives to satisfy these constraints, the Nordmin Group (including Nordmin Engineering Ltd.) (Nordmin) was able to define a concept and design that addressed risks to the project, met production requirements, cost and schedule constraints and improved constructability. INTRODUCTION The Elk Creek Project, owned by NioCorp Developments Ltd. (NioCorp) presently, will become Nebraska’s first underground metals mine, built to access the Elk Creek Niobium/Scandium deposit. The development of the design for this facility required the review of challenging hydrogeological conditions, mine production requirements, demanding CAPEX limitations and tight schedules. Through an extensive review of various alternatives to satisfy these constraints, the Nordmin Group was able to define a concept and design that addressed risks to the project, met production requirements, cost and schedule constraints and improved constructability. This paper will review the challenges faced, and solutions arrived at in a demonstration of a methodology used in designing a modern shaft and hoisting plant. THE CHALLENGES Nordmin became involved in the project in 2015, invited by the Feasibility Study Consultant to provide input on estimating and layouts for the new operation. Many of the fundamental decisions regarding mining methodology, water handling and production strategy had been made, and the scope of Nordmin’s involvement was very limited in this first engagement. Following submission of that study in late 2017/early 2018, Nordmin had opportunity to discuss the overall effort, and its many assumptions and challenges, with NioCorp’s project team. The dialogue quickly identified several significant opportunities to improve the project construction schedule and costs along with the mine production profile and operational approach while reducing project risk."

Jan 1, 2019

By Michael L. Free, Zongliang Zhang, Joshua M. Werner

In the copper electrowinning process, short-circuiting between the anodes and the cathodes is a critical problem that changes the current allocation among local electrodes and reduces the current efficiency significantly. Possible reasons for the short-circuiting include the improper alignment of the electrode, bent electrode, and nodules on the cathode surface. In typical tankhouses, there is often a time between shorting and harvesting as well as between shorting and clearing. Therefore, it is inevitable to have short-circuiting affect the operation performance in practical electrowinning operations. However, one key piece of missing data is the actual measurements of the current efficiency as short-circuiting is approached and allowed to continue. Consequently, short-circuiting tests that allow tracking of electrowinning parameters before and after short-circuiting were performed in a laboratory scale Cu electrowinning cell. In the tests, the electrode spacing change, the mass of electrodeposited Cu, the deposition morphology, and the current allocation as well as the current efficiency before and after short-circuiting were studied. Results show that the short- circuiting affects the current allocation and the current efficiency significantly. Moreover, it changes the Cu deposition on the short-circuited cathode and therefore causes problems for Cu electrowinning. INTRODUCTION Electrowinning (EW) is commonly used for copper production around the world, with a rated production from 120,000 to 200,000 tons per year per operation (Wiechmann, Aqueveque, Henríquez, Muñoz, & Morales, 2014). In Cu electrowinning, Cu ions in the electrolyte are reduced on the cathode surface as metallic copper with a DC current, while on the anode, a counter anodic reaction occurs with oxygen gas generation (Werner, Zeng, Free, Zhang, & Cho, 2018; Zhang, Werner, & Free, 2018; Zhang, Werner, & Free, 2018). In copper electrowinning, an average current density of 200 to 400 A/m2 is applied to the electrowinning cell (Wiechmann et al., 2014). As the Cu deposits on the cathode, the distance between anode and cathode reduces and most of the time, short-circuiting forms between some anodes and cathodes in a large tankhouse containing thousands of electrodes, reducing the current efficiency and increasing the energy consumption significantly. The current efficiency in well-run copper EW operations is typically above 90 %. Most of the unused current is wasted through short circuits, stray currents, and by-byproduct formation (Davenport, King, Schlesinger, & Biswas, 2002; Wiechmann, Vidal, Pagliero, & Gonzalez, 2002; Wiechmann, Morales, & Aqueveque, 2010). Usually, in the tankhouse, the short-circuiting increases the current in the involved cathode by up to 1500 A (Wiechmann et al., 2010). Multiple reasons can cause the short-circuiting in the cell, and the most common reasons include the uneven current distribution among cathodes, tilted electrodes, bent electrodes, and nodules formed due to the large deposition roughness. Recognition of this challenge in the Cu electrowinning process has led researchers to perform many investigations on approaches to avoid and reduce the short-circuiting in the cell. A series of research investigations have been conducted to develop and improve a new type of intercell bar for Cu electrowinning (Wiechmann et al., 2002; Wiechmann, Morales, Aqueveque, & Burgos, 2007; Wiechmann, Morales, & Aqueveque, 2009; Wiechmann et al., 2010; Wiechmann et al., 2014). The associated test results show that this new design limits the occurrence and magnitude of short circuits and allows operation at higher current densities. Sandoval, Clayton, Dominguez, Unger, and Robinson (2010) stated several keys to control short circuits, such as maintaining the straightness of the cathode, removing “edge hair” from the cathode starter, and maintaining electrode alignment in the cell. Prengaman and Siegmund (1999) from RSR technologies developed a new Pb-Ca-Sn anode to prevent the anode deformation, thereby reducing the short-circuiting in the Cu electrowinning cell. Although the negative effects of short-circuiting on the Cu electrowinning have been recognized, and various investigations have been conducted to address this issue, the actual measurements of the electrowinning behavior and performance as short-circuiting is approached and allowed to continue are seldom reported. However, in typical tankhouses, there is a time between when shorting occurs and harvesting/clearing are performed. One key factor in determining the harvest time is the frequency of shorting. Consequently, we have performed short-circuiting tests that allow us to track electrowinning parameters before and after short-circuiting.

Jan 1, 2019