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By T. Patching

"MINING operations at the Star Key Mine provided an example of a small-scale prairie-type coal mine that employed several interesting procedures and achieved good productivity with low capital expenditures. The mine was located about 11 miles north of the city of Edmonton, on the north side of the Sturgeon River valley. It opened in 1945 and operated until 1979. Output in the winters reached about 500 tons a day in the late 1940s, but in the spring and summer periods activity was usually limited to only servicing and development work. In the later years of the mine's life sales fell to less than 100 tons per day. Most of its output was bought for domestic use in furnaces and stoves so production of large sizes was desired. A modest market for stoker coal did exist for a time. Screenings (-5 / 8 in.) were sold to the city's power plant until it converted to natural gas and were also used for thawing frozen ground for work on pipelines. These markets were comparatively small and the small sizes did not command as high a price as did the larger sizes. However the extension of natural gas services to rural markets cut into the domestic heating market and finally led to closure of the mine. The coal was strong when freshly mined, but because of its high inherent moisture content, about 25%, it did not store well on surface but tended to disintegrate. It was of subbituminous Crank with a heating value of about 8000 Btu per pound and burned with a clean flame. The seam mined was in the Edmonton Formation, of Upper Cretaceous age, at a depth of 60 to 80 ft below the surface. The seam in the mine area was about 11 feet in thickness but thinned gradually to the north and south and split to the west (1). It was approximately horizontal but gentle rolls in the floor of several feet in height necessitated brushing in places to produce suitable haulage grades. The floor material was of bone and clay, overlying a thin layer of coal. Above the seam there were only a few feet of mudstones and weak sandy shales while the remainder of the overburden was clay and unconsolidated glacial till and surface soil. Because of the soft floor and weak roof material the miners usually left about a foot of coal in the floor to provide a firm working surface and about a foot of coal above the rooms and two feet of coal above the entries to provide a competent roof."

Jan 1, 1985

By Li Qihou, Liu Zhihong, Liu Zhiyong, Yao Wei

Black copper sludge, which mainly contains copper, arsenic and a little antimony, bismuth, selenium, tellurium, is an important recyclable secondary resource. In this paper, selective removal of arsenic from black copper sludge by alkaline oxidative pressure leaching was studied. The influences of alkali concentration, leaching time, leaching temperature, oxygen pressure, liquid-solid ratio and other variables on the leaching rate of arsenic, copper and tellurium were investigated. The results were as follows: when the leaching process was conducted at initial alkali concentration of 1mol/L, liquid-solid ratio of 10, stirring rate of 500r/min, leaching temperature of 140°C for 5 h and oxygen pressure of 2.5MPa, the leaching rates of arsenic and tellurium were 85.43% and 59.7% respectively, and copper, antimony, bismuth were not leached. Hence, arsenic has been removed effectively from black copper sludge, while the valuable metals such as copper, antimony and bismuth remained in the residue that can be further recovered easily. INTRODUCTION Black copper sludge is a byproduct in the purification process of copper electrolyte. Arsenic, antimony and bismuth are major harmful impurities in copper electrolytic refining system. The accumulation of these impurities will cause adverse effects, such as the increase of electrolyte resistance and power consumption and deterioration of cathode copper quality (Wang, Liu, & Zhang, 2001). Hence, electrolyte purification is an essential part of the electrolysis process (Liang, Chen, & Guo, 2009). Electrolyte purification methods include electrowinning (Davenport, King, Schlesinger, & Biswas, 2002), solvent extraction (Navarro, Simpson, & Alguacil, 1999), ion exchange (Arroyo et al., 2017), coprecipitation (Wang et al., 2003, 2018; Xiao et al., 2007; Xiao, Cao, Mao, Shen, & Ren, 2014), activated carbon adsorption (Navarro & Alguacil, 2002), electrodialysis (Cifuentes et al., 2015), redox process (Zhou, Peng, Zheng, Ma, & Cui, 2011; Peng, Zheng, Zhou, & Chen, 2012; Zheng, Peng, Ke, & Chen, 2013), among which the electrowinning method is simple in operation and effective in impurities removal, making it the main process for current electrolyte purification. Some copper ions can react with harmful impurities to form complex black copper sludge by controlling the operation conditions of electrowinning process. Based on the estimated data provide by Davenport et al. (2002), an annual output of 400,000 tons cathode copper could produce 432-2168 tons of black copper sludge by electrowinning method. Black copper sludge contains 30~50% copper, 15~30% arsenic and a small amount of rare elements such as selenium and tellurium. Thus, recovery of valuable metals can bring huge economic interests. Moreover, arsenic in these sludges could cause significant environmental pollution. Therefore, effective removal of arsenic from these sludges is urgent and significant.

Jan 1, 2019

By J. Lu, D. B. Dreisinger, G. Houlachi

"Zinc metal is primarily produced by electrowinning from sulfate solution. Zinc electrowinning is very sensitive to impurities such as Co and Ni. These impurities have to be removed to produce high-quality zinc and maintain a high zinc current efficiency. Zinc dusts are used to remove these impurities by cementation. However cobalt cementation is extremely slow and it is typically activated by the addition of Cu and Sb. Even with the use of Cu-Sb activation, cobalt is still difficult to remove. The zinc dust used for cobalt cementation is negligible compared to that consumed for hydrogen evolution. In some plants up to 6-7% of electrolytic zinc is used for cementation. This is a great waste of energy. Cu-Sb activation is not well understood. The objective of this study was to further understand the Cu-Sb activation and then find a way to reduce hydrogen evolution and promote cobalt cementation. The investigation reveals how hydrogen evolution (or zinc dissolution) and cobalt cementation and re-dissolution change with temperature, zinc dust dosage, the concentrations of copper and antimony and time. In the Cu-Sb activation, copper is deposited on a zinc particle to produce a porous shell. The zinc particle as an anode dissolves while the shell as a cathode provides a place where cobalt cementation and hydrogen evolution occur and they are catalyzed by antimony. A certain level of antimony (around 0.1 mg/L) is a key to maintain the catalytic effect of the shell. At a higher antimony concentration, cobalt cementation is accelerated slightly while hydrogen evolution is increased greatly. When a zinc particle dissolves to a certain degree, cobalt loses its cathodic protection and starts to re-dissolve. The change in cobalt concentration is determined by the rates of cobalt cementation and re-dissolution. The operation conditions can be optimized to maximize the cobalt removal and minimize the zinc dust dosage.INTRODUCTIONMost of the world’s zinc metal is produced by electrowinning from sulfate media. Many impurities such as iron, copper, cobalt and nickel are co-extracted into solution with zinc. Iron, mercury, indium, lead and tin are removed by co-precipitation as hydroxides. Copper, nickel, cobalt, aluminum, arsenic and germanium are partially co-precipitated with iron while cadmium, magnesium and chloride remain in solution. Cobalt, copper, cadmium, nickel and etc. are subsequently removed by cementation with zinc dust to achieve an acceptable current efficiency and high purity metal product for the economical production of zinc. Zinc provides a strong thermodynamic driving force for the cobalt cementation:"

Jan 1, 2016

By Daniel Hodouin, Mélanie Proulx, Claude Bazin

"Collector and frother are commonly used to induce the floatability of valuable minerals in flotation plants. These reagents are usually distributed down a flotation bank or sometimes added in a single location at the top of the bank. A review of the literature and some experimental results obtained from laboratory and full scale trials indicate that the way flotation reagents are distributed down a flotation bank may influence the flotation response of coarse particles. The paper summarizes these observations which show that the distribution of reagents is a potential tuning variable for process optimization.IntroductionFlotation is in many cases the only way to concentrate sulphide and some oxide minerals. Flotation is a concentration process that uses particle surface composition to induce the selectivity by mean of hydrophobic reagents. The influence of particle size on flotation was firstly analyzed in the 50's1·2 and intensively investigated in the 70's, early 80's and 90's3-18. Most of the observed results show the existence of an optimum flotation size range9 whose recovery is not significantly influenced by a reduction of collector dosage""18. This result brought the idea of using a small portion of the total collector dosage at the top of a rougher bank and to distribute the remaining through the end of the bank to ensure sufficient collection of the coarse particles that need more hydrophobic coverage to attach to air bubbles and be lifted to the surface.This paper discusses the idea of using the distribution of reagents down a flotation bank to tune the operation of the process. Unlike the literature on the effect of reagents (collectors, frother...) dosage that is dense, little attention has been put on the distribution of reagents in a flotation bank. Although research on this subject has not been pushed very far, some operators have recognized the potential of a strategic addition of reagents down a flotation bank19. Except for a few exceptions, the proportions of the reagents to be distributed down a flotation bank are decided according to operator's experience instead of process analysis in relationships with particle size. This paper recalls observations made by different researchers that lead to the idea of using the distribution of reagents as a tuning variable for the flotation process. The paper does not present a detailed analysis or optimization scheme for tuning the distribution of reagents, but analyzes some plant and laboratory results with a view toward the development of a testing methodology and ultimately a validation or rejection of the idea."

Jan 1, 1998

By D. R. Hatton

To obtain sufficient material on which to perform copper-molybdenum separation testwork a variability set of drill hole reject samples amounting to several tons was put together to include the different ore types proportions expected in the copper circuit. The samples were treated in a copper circuit pilot plant and enough mass of bulk concentrate was recovered to conduct copper moly separation tests in the laboratory. Tests consisted of stage flotation reactor (SFR) rougher pilot plant tests, conventional laboratory rougher, cleaner and 4 stage open circuit cleaner tests. The mineralogical liberation analysis of the feed and concentrate was used to model the effect of liberation on the grade-recovery relationship of the different samples. Mineral kinetic parameters for the various samples were determined by calibration to the test data. The copper concentrate results as simulated from previous studies on the same deposit were used to simulate the variability in head grade and throughput expected for the feed to the copper molybdenum separation circuit for each of the first 10 years of production. This variation in feed was then simulated through the separation circuit to obtain optimised molybdenum grade recovery curves to forecast the circuit production for each of the first 10 years. INTRODUCTION Incorporating Geometallurgy into the design for a copper molybdenum separation is especially challenging due to the low mass of final concentrate typically produced during testwork. This is due to the low molybdenum head grades and high overall molybdenum upgrading expected. With molybdenum head grades of around 300 ppm, 1 tonne of copper circuit feed is expected to produce around 300 g of molybdenum concentrate in the plant. In the laboratory however, this may end up only resulting in a few grams. Obtaining sufficient mass for laboratory testing therefore requires starting with a large initial sample and careful testwork planning to maximise the utilisation of this mass. A testwork program was designed to investigate the effect of ore variability on the copper molybdenum separation circuit (here after referred to as the molybdenum circuit). Composite samples were selected to represent the various ore types, copper content, moly content, pyrite content and clay contents expected in the copper circuit feed. These samples were run through a bulk copper pilot plant to create the feed for the molybdenum testwork. The molybdenum testwork consisted of rougher kinetic flotation tests, rougher cleaner kinetic tests, rougher cleaner kinetic tests with regrind, stage flotation reactor (SFR) pilot plant tests, open circuit cleaner tests. These tests were used to calibrate the ore kinetics in the IGS-FLEET model for simulation of the molybdenum circuit performance.

Jan 1, 2019

By S. Onuki, T. Ando, M. Tejima, R. Enomoto, Y. Okada

"Al–Si–Cu–Mg die casting alloys are widely used in the automotive industry, and recent advancements in high pressure die casting technology enable heat treating products to achieve desirable properties. However, CO2emissions associated with die-casting production lines should be reduced to address the global warming issue. In this study, the relation between the microstructures and thermal stability of precipitates formed during isothermal aging at 423, 485, and 558 K of Al-11.0 mass%Si-1.8 mass%Cu-0.2 mass% Mg (JIS ADC12) was investigated using Vickers hardness tests and transmission electron microscopy (TEM). The hardness tests and TEM results showed that Si compounds and Al–Cu compounds enhanced the mechanical properties of the alloy and that the predominant compound affecting the mechanical properties differed between specimens treated at 423 K and those treated at 558 K. Moreover, additional aging at 423–458 K resulted in a material with a Vickers hardness higher than that of the isothermally single-aged material. Differential scanning calorimetry, differential elemental analysis, and selected-area diffraction analysis using TEM suggested that the combination of Al–Cu compounds precipitated at low aging temperatures and Si compounds precipitated during preliminary aging increased the mechanical strength, resulting in a maximum hardness of 130 HV.INTRODUCTIONAl–Si–Cu–Mg die casting alloys have well-balanced mechanical properties, machinability, and cast ability. These alloys are therefore widely used in the automotive industry because of their superior characteristics, accounting for>90% of the aluminum used in die casting production. As a substitute for iron-based materials, these alloys have contributed to the development of lightweight vehicles.Dong et al. (2014) reported that the heat treatment of die casting materials is effective for changing their dimensions under high temperatures as a result of the precipitation of solid-solution elements. Recent advancements in high pressure die casting technology have led to the application of heat treatments of Al–Si–Cu–Mg alloy products to reduce fuel consumption and satisfy customer needs. Although attempts to reduce CO2emissions have focused on fuel consumption in cars and fuel efficiency regulations have become increasingly strict in recent years, CO2discharged from production lines should also be reduced without adversely affecting the properties of the product. Heat treatments for artificial aging after solution treatment (T6 treatment) and after casting (T5 treatment) are CO2discharge sources, which together with a holding furnace account for >80 % of the CO2discharged during the casting process."

Jan 1, 2018

By C. W. Chao, M. Melashvili, F. Hilário

The processing of a bulk concentrate containing copper as chalcocite, enargite, bornite and chalcopyrite along with pyrite and precious metals of silver and gold was studied. A leaching study was undertaken to examine the application of the atmospheric acid leaching process to a bulk copper concentrate for copper recovery followed by cyanidation or thiosulfate leaching for gold and silver recovery. The extraction of copper exceeded 90% under some conditions and the recovery of gold and silver reached a maximum of 44 and 45%, respectively. A sample of bulk concentrate was roasted in air to oxidize the sulfides with release of sulfur dioxide followed by acid leaching of the roaster calcine and cyanidation of the acid leach solution. The copper extraction reached over 92% and the gold extraction exceeded 88%. The sulfur dioxide may be captured for acid production. INTRODUCTION The processing of a large primary copper deposit in Indonesia is under investigation. One of the possible routes for processing involves flotation of a bulk copper-gold concentrate followed by treatment by an atmospheric leach/SX-EW process for copper recovery. The atmospheric leach residue would then be leached for recovery of gold and silver. A second route involves roasting of the concentrate followed by acid leaching and SX-EW for copper recovery. The leach residue from the roaster calcine would then be leached for recovery of gold and silver. The roaster off-gases could be used for sulfuric acid manufacture, enabling the production of a valuable by-product. The Cobre Las Cruces (CLC) process (Munoz, Garcia, & Delgado, 2017) was designed to treat a resource of copper discovered close to Seville, Spain. The process was developed by Outotec with MK Gold (original project owner). INMET Mining subsequently purchased the deposit from MK Gold and built the copper processing facility. The CLC process is similar to the Mt. Gordon Copper process; except that OKTOP reactors have been installed instead of a low temperature, low pressure autoclave process. The plant was designed to treat 1.3 mtpa of ore at a grade of 6.2% Cu to produce 72,000 tpa of cathode copper over a 15 year project life. The main gangue mineral at Cobre Las Cruces was pyrite. The Cobre Las Cruces ore is therefore similar to a rougher flotation concentrate containing copper minerals and pyrite. The roasting of a rougher flotation concentrate can be designed to produce hematite, basic copper oxides and sulfur dioxide off gas. The acid leaching of the basic copper oxides may produce soluble copper for SX-EW recovery of copper. The sulfur dioxide off-gas may be used to make sulfuric acid.

Jan 1, 2019

By R. Saulters, M. Garshasb

The Transportable Moisture Limit (TML) is the copper concentrate moisture content beyond which liquefaction may occur during shipping. This study indicates that TML is a function of mineralogical characteristics of concentrate, which in turn may directly affect moisture retention properties. A comprehensive mineralogical and chemical analysis was conducted to determine the impact of composition and mineral content on concentrate’s TML. The analysis indicates that the high TML value is associated with particular minerals such as gypsum, swelling clay, muscovite and pyrite. Since higher TML values allow higher moisture contents for shipping, this property is desirable for concentrate inventory management and shipping flexibility. Scarcity of minerals that may increase TML contributes to limitations of shippable inventory. INTRODUCTION TML is the maximum moisture content of copper concentrate beyond which liquefaction may occur. Moisture content above the TML is considered a shipping hazard as it may result in loss of lives and cargo. The details of TML testing are given in ISO 12742:2007 (ISO 12742). In principle, TML is 90% of the Flow Moisture Point. The method of testing for Flow Moisture Point and TML is based on procedures that provide pulsations that simulate the cyclic force of ocean waves to determine moisture levels at which the onset of concentrate flow may occur. International Maritime Organization (IMO) provides guidelines for developing and approving TML testing and moisture determination in the context of its Moisture Management Plan (MMP) for shipment of copper concentrate cargoes (IMO 2013). The IMO regulations are specifically provided in MSC.1/Circ. 1454 “Guidelines for developing and approving procedures for sampling, testing and controlling the moisture content for solid bulk cargoes which may liquefy”. Since each port has its own operational characteristics, resource availability, peculiarities of governmental structure, and desired methods of interaction with the authorities, it is necessary to adopt the IMO guidelines to the local requirements. In order to establish a TML and MMP measurement system, it is required to implement and follow proper guidelines for the relevant procedures. For example, sampling procedures are developed in accordance of ISO 12743, “Copper, lead, zinc, and nickel concentrates – Sampling procedures for determination of metal and moisture content”. Moisture content of copper concentrate samples are determined in accordance with ISO 10251, “Copper, lead, zinc, and nickel concentrates – Determination of mass loss of bulk material on drying”. To assure strict adherence to the MMP procedures, the overall system needs to be integrated into the site QA/QC system, which would make them auditable and certified in accordance with standards such as ISO 9001, ISO 17025, and ISO 17020.

Jan 1, 2019

By G. Morin, G. Marin, M. C. Parr, J. Carignan

"Investment casting offers many advantages for the near net shape production of complex geometries, however, molten titanium reacts with most of the refractories used in investment casting shelling systems, which leads to the formation of an oxygen contaminated surface layer. This layer, called alpha case, is 0.1 to 0.6 millimeters in thickness according to the casting section size and must be removed either by mechanical or chemical milling processes. Alpha case removal processes affect casting tolerances and may contaminate the part intended for medical applications (e.g. embedding of abrasive grit media in the alloy). A shelling system based on an yttria prime slurry has been developed and was found to almost suppress the mold metal reaction. Yttria is an expensive refractory, but slurry life was found acceptable for small castings using a closed container and flow coating process instead of conventional pattern dipping. Alpha case depth has been determined for Ti6Al4V by metallographic techniques and microhardness measurements on different step block casting cross sections. Alumina and yttria second prime stuccos were evaluated in this study. The shelling system minimizes finishing operations in the production of casting for medical applications.INTRODUCTIONTitanium investment castings have been in the commercial market since the beginning of the titanium era in the mid 50’s. Investment casting is a cost effective way to produce complex parts for medical, aerospace and chemical processes applications. Investment casting of titanium is difficult due to the high reactivity of titanium with air and refractories used for shell mold building especially for the face coat. Melting is done under vacuum and no shelling system has been proven effective to suppress the mold-metal reaction yet. The main issue with the mold-metal reaction is the formation of an oxygen enriched zone at the surface of the castings named alpha case (Frueh et al., 1997; Kim, 2002; Saha, 1986). The alpha case is brittle, reduce the fatigue life in service of the parts and must be removed by chemical milling.For industrial purposes, calcia partially stabilized zirconia based slurries have become widely used in titanium investment casting for primary coats due to a combination of limited metal shell reaction. Calcia stabilized zirconia slurries have an acceptable shelf life and offer an interesting production window. Our experience shows that a resulting alpha case depth is 280 micrometers for a casting section of 6.35 mm in thickness is typical with such a shelling system (Morin, 2013)."

Jan 1, 2016

By Mory M. Ghomshei

"The cost of deep underground mine cooling depends on a number of geological, environmental and economic factors such as the temperature gradient in the mine, the outside air temperature, and the local price of energy. While geological factors, being static, can be controlled to some extent during the mine design, the environmental and economic factors dynamically change and therefore can only be controlled through a flexible mine ventilation system.With an average temperature gradient of 20–25°C/km, the virgin rock temperature at 2,000 m can reach 45–60°C. Conventional cooling at such depths could lead to exorbitantly high costs. In-mine geothermal heat pumps can considerably reduce the cost of cooling through energy communication with the centralized absorption chillers, which service the mine from the surface.It is shown that in-mine heat pump systems can be a viable option for reducing operating expenses and providing some degree of flexibility to cope with fluctuating energy costs. The price of Natural gas and electrify are the most important variables for optimizing the level of interaction between the surface and inmine cooling systems. Deep mine waters can be used to carry a portion of the mine heat to the surface, where the heat from mine can be used as a supplementary energy source for absorption chillers. As an extra benefit, the heat output of the chillers can be used for district heating and or greenhouse growing.The case of cooling deep levels of deep mines in Sudbury Canada (e.g. Creighton Mine) has been discussed. It is shown that an initial 1MW in-mine geothermal heat pump system can save significantly on the cost of cooling. Such a demonstration project will allow R&D on viability of larger similar systems to be incrementally added in the future.INTRODUCTIONThe amount of heat generated in deep and ultra-deep mines depend primarily on the regional temperature gradient and the thermal conductivity of mine rocks. The conductive heat flow is quantifiable and can be minimized or optimized in the process of mine design. Another important heat source can be related to underground equipment such as diesel machinery which often have low efficiencies (approximately 25–35%). This heat source can be controlled and reduced though improving efficiencies or using electrical equipment, which offer a much better efficiency (approximately 80–90%). Mine waters can also be a significant heat source (convectively bringing large amounts of heat into the mine workings)."

Jan 1, 2019

By George J. Cymerman

Since 1978, Syncrude Canada Ltd. has been expanding steadily its oil sand mining and processing capacity while reducing unit energy requirements. At the current rate of ~180 M tonne/year oil sand, the originally adopted Karl Clark?s 80°C version of the Hot Water process would require some 53 PJ (53 x 1015 Joule) heat. If this amount of heat were to be derived from burning natural gas it would cost $369 M/y (at $7/GJ). Fortunately, effective recovery of waste heat from the upgrading operations at the Mildred Lake site obviated the need for water heating with natural gas. The oil sand hydrotransport technology, first utilized commercially in the mid-nineties at Syncrude?s North Mine, reduced slurry temperature from ~80°C to ~50°C, cutting the unit heat requirement from 293 MJ/tonne to 183 MJ/t. However, in the new 100 Mt/year Aurora Mine even a 50°C process temperature would still require 18 PJ of heat energy per year. Only limited quantity of 65°C water could be imported from Mildred Lake site and some waste heat would be available from Gas Turbine Generators. A new, Low Energy Extraction (LEE) process was developed to minimize heat requirements by further reducing process temperatures to 25°C in the slurry pipeline and 35°C in the Primary Separation Vessels (PSV). Unfortunately, when applied to certain ore facies in Aurora, the LEE process yielded disappointing bitumen recoveries, necessitating an increase in slurry temperatures from 25°C to 35°C in the pipeline and from 35°C to 42°C in Primary Separation. In 2004, an Optimized Low Energy process (OLE) was developed and piloted at Syncrude Research using a 2000 kg/hour pilot plant. In the OLE process, most available heat is applied up-front to produce 45°C pipeline slurry at a density of 1.58 kg/l. Once the pipeline conditioning process is completed, the slurry is flooded with cold (10°C) water, resulting in PSV temperature of 34°C. The overall heat and water requirements are the same, or slightly lower, compared to those of the LEE process ? but without the penalty in bitumen recovery. Processing 100 Mt oil sand at 34°C (96 MJ/t), rather than at 50°C (183 MJ/t), will save $61 M/y (@ $7/GJ) and reduce green house gas emissions by ~47 %. The OLE process is currently being implemented at the Syncrude Aurora Mine.

May 1, 2006

By J. Bouchard, M. Levesque, D. Georges-Filteau, P. Radziszewski, G. LeBlanc

"Grinding mills are infamous for their extremely low energy efficiency. This paper reports the results of an investigation that characterized energy outputs in industrial grinding mill circuits. It identifies and discusses the potential for recovering energy from grinding circuits. Agnico-Eagle Goldex Division, Mine Canadian Malartic, and New Gold New Afton Mine participated in the study by providing operating data for three semi-autogenous grinding mills and four ball mills. Results showed that on average, 79% of the supplied electrical energy was converted to heat absorbed by the slurry, 8% was lost through the drive system, and approximately 2% was transmitted to the ambient air. Only 9% of the input energy was actually used for grinding. The remaining 91%—wasted as heat—could potentially be recovered using suitable technologies or integrated energy management systems. This topic remains to be further investigated. ¦ KEYWORDS Comminution processes, Energy recovery potential, Grinding mills RÉSUMÉ Les broyeurs sont tristement célèbres pour leur rendement énergétique médiocre. Cet article présente les résultats d’une étude qui caractérise les rendements énergétiques des circuits industriels de broyage et examine les possibilités de récupération d’énergie dans les circuits de broyage. La mine Goldex d’Agnico Eagle, Mine Canadian Malartic et la mine New Afton de New Gold ont participé à l’étude en fournissant des données de fonctionnement pour trois broyeurs semi-autogènes et quatre broyeurs à boulets. Les résultats montrent qu’en moyenne 79 % de l’énergie électrique fournie sont convertis en chaleur absorbée par la pulpe, 8 % sont perdus dans le système d’entraînement et environ 2 % sont dispersés dans l’air ambiant. Seuls 9 % de l’énergie primaire sont, de fait, utilisés pour le broyage ; les 91 % restants (perdus sous forme de chaleur) peuvent potentiellement être récupérés à l’aide de technologies adaptées ou intégrées dans des systèmes de gestion de l’énergie. Cette question demeure toutefois ouverte. ¦ MOTS CLÉS broyeurs, potentiel de récupération de l’énergie, procédés de comminution INTRODUCTION Grinding is largely recognized as a very inefficient process; energy efficiency estimates range from < 1 to 2% (Fuerstenau & Abouzeid, 2002; Tromans & Meech, 2002, 2004) when comparing the input energy to that required for generating new mineral surfaces. Criticizing an “ill definition of the reference for the output energy,” Fuerstenau and Abouzeid (2002) proposed a “more meaningful baseline” would be “the energy for producing new surface area by the ¬mens.” Based on this reference, they concluded that “the ball mill is reasonably efficient energetically,” exhibiting an efficiency of approximately 15% for quartz."

Jan 1, 2019

By M. Cai

"It is well known that Peak Particle Velocity (PPV) plays a critical role in triggering rockburst damage in burst-prone underground mines. To estimate the PPV distribution around excavations, advanced numerical modeling can be used. In wave propagation numerical modeling, velocity field is often assumed to be uniform around an excavation. However, it is observed that wave velocities near mine openings are not constant, but rather non-uniform. Wave velocity in a rock mass depends on many factors such as rock quality and confinement. In particular, confinement has a large influence on wave velocity. The confinement around an excavation is low at the wall boundary and increases gradually and reaches a constant value far away from the excavation boundary. Hence, stress redistribution due to excavation will have a profound influence on wave propagation in underground mines. A better understanding of the influence of excavation effect on wave propagation helps to interpret the field observation of rockburst damage and improve mine safety. For this purpose, a non-uniform velocity model need to be constructed and the influence excavation effect should be taken into account in numerical wave propagation modeling.In the present study, a non-linear velocity model is developed firstly by considering the influence of confinement on wave velocity. This model links wave velocity to confinement explicitly. A coupled numerical method is then developed and used to study the seismic wave pattern around excavations in underground mines. Two codes, FLAC and SPECFEM2D, are coupled in the analysis. The motivation of developing the FLAC/SPECFEM2D coupled approach is to take merit of each code, such as the stress analysis capability in FLAC and the powerful wave propagation analysis capability in SPECFEM2D. The coupling process is realized through three steps: (1) static stress redistribution around an excavation is analyzed in FLAC; (2) data for coupling analysis, such as non-uniform velocity field (based on the new developed velocity model), mesh, absorbing boundaries, etc., are generated by FISH functions in FLAC and passed to SPECFEM2D; and (3) wave propagation modeling is performed in the SPECFEM2D considering a fault-slip seismic source. Because stress redistribution around the excavation is considered, non-uniform velocity field for the SPECFEM2D model can be obtained and this is a notable feature of this study. Very large difference in the S-wave response in these two modeling approaches with uniform and non-uniform velocity fields can be seen from the modeling results.Using the FLAC/SPECFEM2D coupled modeling approach, accurate velocity models can be constructed and this in turn will help to produce accurate PPV distribution around the excavations. Accurate PPV contours can be used to identify high rockburst risk zones for improved dynamic ground support strategy."

Jan 1, 2015

By K. E. Waters, J. E. Nesset, F. Rosenblum, J. A. Finch, S. Moon

"Some materials when exposed to various conditions can exhibit a rise in temperature without requiring an external heat source. Generally referred to as pyrophoric or self-heating substances, they challenge the safety of materials handling, storage and transportation. Reliably assessing the self-heating potential of materials is thus essential. Over the years a number of test protocols have been developed. A widely accepted method is that recommended by the United Nations to certify the safety of shipping pyrophoric goods. Our test work on sulphide materials has shown that the U.N. test can give false negatives if the sample oxidation history is not considered. Depending on the degree of oxidation (“weathering”), the same sulphide material can exhibit self-heating ranging from safe to hazardous. For concentrates, oxidation can occur during storage and transport making the sample weathering history an important consideration in the test work. To make the U.N. test, and other single-stage self-heating procedures, more reliable in assessing self-heating potential when testing sulphide materials, it is suggested that a weathering step be incorporated in the test protocol. This test work for two sulphide samples, an ore and a concentrate, demonstrates the impact of incorporating a weathering stage in the test protocol. INTRODUCTIONSelf-heating occurs when exothermic reactions produce heat faster than can be dissipated to the surroundings. Numerous substances self-heat, ranging from wood chips and powdered milk to coal and sulphide minerals (Beever & Crowhurst, 1989; Hudak, 2002). Mixtures of sulphide minerals, most notably those containing the iron sulphide pyrrhotite (Fe(1-x)S), commonly encountered in the extraction of base metals (e.g., copper, zinc, lead, and nickel), are particularly prone to self-heating (Rosenblum, Nesset, & Spira, 2001; Payant, Rosenblum, Nesset, & Finch, 2012). In the mining and subsequent processing stages, the minerals are progressively liberated and finely ground, then separated, usually by flotation, with the resulting concentrates stored and transported to smelters for metal extraction. At most of these stages the sulphides are exposed to both moisture and oxygen, conditions that promote partial sulphide oxidation and the formation of elemental sulphur. Studies have shown that pyrrhotite, in particular, readily oxidizes to form elemental sulphur (Bernier & Li, 2003; Belzile, Chen, Cai, & Li, 2004; Somot & Finch, 2010). The elemental sulphur becomes fuel for further self-heating as the temperature of the mineral mixture increases above 100 oC (Rosenblum & Spira, 1995; Somot & Finch, 2010). It is believed the propensity of pyrrhotite to self-heat is a result of its iron-deficient, and hence somewhat unstable structure. It is worth noting that pure metal sulphides (with the exception of pyrrhotite) do not exhibit appreciable self-heating. A mixture of minerals sufficiently different in their electrochemical rest potential is required to initiate self-heating at ambient temperatures (Payant et al., 2012)."

Jan 1, 2016

By R. Kong, X. W. Zhang

"High pressure (up to 50 MPa) isotropic compression tests were performed on marble, gneiss and sandstone specimens, of which represents the typical rocks, to investigate isotropic compression deformation characterization at room temperature. Experimental data with respect to volume change in loading and unloading are presented. An apparent volume change inflection point (sensitive stress) was observed at a lower pressure range for porous sandstone, indicating that the elastic-plastic behaviour arising in the isotropic compression stage, but for tight crystalline rock, the sensitive stress is much higher, not probed prior to 50 MPa. The irrecoverable volumetric strain in the unloading can be seen for sandstone as well. The anisotropic deformation was discussed in terms of the axial and radial strain measurement. It was found that essential features of mechanical anisotropy are the irreversible in deformation range. The deformation feature in the isotropic compression tests is a very complex issue due to the factors concerning the stress history, unloading in sampling and in-situ stress. INTRODUCTIONThe isotropic or hydrostatic compression condition that stress is compressive and has the same magnitude for all direction, which is expressed as s1=s2=s3 in principal stress space. A yield locus of soils subjected to the isotropic and anisotropic compression was widely known, for instance Cam Clay model, which forms a closure cap model. However, the similar isotropic compression yield curves for rocks, especially for hard rock, was not established due to the geo-stress in the rock mass is much less than the socalled isotropic yield stress. And hence, the previous experimental studies on the hard rock have been made tend to on the stress-strain-strength relationship (Handin & Hager, 1957; Edmond & Paterson, 1972; Menéndez et al., 1996). Perhaps the yield curves related to an increasing in the hydrostatic pressure is not forced to construct for hard rocks, undoubtedly, the deformation data in this processing is valuable to assess the bulk module (Fabre & Gustkiewicz, 1997), anisotropy (Niandou et al., 1997), field stress (Hakala et al., 2007). Furthermore, many rocks store much elastic volume strain energy link to the hydrostatic pressure exerted, following a deviator stress applying, to drive the unstable brittle fracture for hard rock. Latter was studied to interpret the incentive of rock burst. However, volume change characteristic in the isotropic compression behaviour have been scarcely studied experimentally to our knowledge, less application related to the fracture in hard rock as well. The mechanical behaviour in the isotropic compression of hard rock is notably affected by the original structure anisotropy, stress history and stress condition before sampling, all of which need reasonably assess. It is also a useful understanding to study isotropic compression that elastic and plastic volume change in the isotropic loading and unloading, and the micro crack evolution throughout it."

Jan 1, 2015

By T. M. Siew

"This paper begins with a brief discussion of the geology of the Newfoundland Zinc Mine area. The results of groundwater studies and pumping tests are then presented. It was decided that groundwater in the mine could best be controlled by conventional sump pumping and grouting. The system is fully described, as well as plans for improvements on the method. LocationThe Newfoundland Zinc Mine is located in the Great Northern Peninsula of Newfoundland, six miles northeast of Daniel&apos;s Harbour. The mine is situated at an elevation of approximately 350 feet above sea level, 5 miles farther inland , west of the long range of mountains which rise abruptly with elevations of over 2,000 feet above sea level.GeologyThe ore deposits occur in lower Ordovician carbonate rocks that form part of the stable St. Lawrence platform, which is mainly composed of (1) the lower Cambrian Labrador group, (2) the upper Cambrian to lower Ordovician St. George group, (3) the disconformity overlying Middle Ordovician limestones and shales of the Table Head formation and, (4) two overlying klippen composed of Cambrian and lower Ordovician rocks.The ore deposit is a Mississippi Valley type of the Tennessee variety, having proven are reserves of 5.3 million tons averaging 9.70/0 zinc from eleven ore zones which exist in an area of 5.5 square miles.Among these ore zones, the L zone is the major underground mining area, and is by far the largest of the known deposits. It has proven are reserves of 3.4 million tons with 11.5% zinc; this underground development has been inprogress since 1974.Stratigraphically, the orebody, which occurs about 300 feet below the disconformities between the Table Head and St. George formations, is situated in the upper St. George formationof lower Ordovician age.The upper St. George formation of the L zone is divided into three main units (Fig . I) : the underlying lower limestone the middle grey dolomite pseudo breccia unit and the overlying siliceous dolomite unit. In this area of the mine, the beds of the St. George formation dip gently west to northwest to 10 to 15 degrees. The orebody, which is generally strata controlled with a southwesterly plunge and which has the form of thin, narrow, but elongate, en-echelon ore bed s, is contained in grey dolomite pseudo breccia. Its dimensions vary from 50 to 900 feet in width and 8 to 120 feet in thickness, and the over-alllength exceed s 6,000 feet.The surficial deposit of the area is a till comprising a heterogeneous mixture of sand, gravel, cobbles and boulders, with an average depth of 20 feet."

Jan 1, 1980

Building block cave operations involves significant commitments to capital and time before arriving at completion or full scale production. The focus of the mine planners is habitually to minimize the upfront capital and achieve target development rates and construction dates. However, management of the construction of block cave projects requires understanding of the strategic/tactical decisions that the project planners can take to increase the value of the project and cope with schedule risks. As block cave projects involve development of vertical and lateral access to the underground footprint and ore body, more attention to uncertainties is required as most of those projects need to confront delays leading to a decrease in value. The standard process of development and construction scheduling is often not planned with provisions to realise uncertainties in the construction phase. Uncertainties and risks such as poor ground conditions and equipment breakdowns are the primary reasons for schedule slippage. As a result most of these projects experience delays. In order to mitigate risk to the project schedule, we centre our attention on increasing the value of the project by spending more money upfront and improving the schedule rather than focus on minimizing the upfront capital. Scenario-based simulations along with project appraisal methods hold great promise to address those challenges and enable mine planners to make better decision and enhancing project value. This paper will present a hybrid simulation based approach to model the process of development and construction based on discrete event (DES) and continuous based simulations (CBS). Such simulations will be conducted for scheduling multiple development headings (Apex, undercut, extraction, and haulage levels). Two scenarios are utilized to investigate the impact of risks on schedule and how flexibility in construction can be incorporated to minimize delays: (1) The bench mark scenario is created based on actual data from an existing case study. (2) The flexible scenario is an alternative way to use primary production infrastructure (Ore Handling System) including conveyer drifts, conveyer belt, and a gyratory crusher in the construction process and that requires the system to be in place before starting lateral development in the footprint. As a result, a truck-load system versus a conveying system will be studied to observe the gain in construction speed and to estimate the premium cost for mucking rock from headings and drawbells and transporting it to surface. Such strategic flexibility would be part of the original budget to make the tactical flexibility available during construction. Continuous cash flow models are used to evaluate and compare the scenarios. This study is an example of how new techniques in planning can be used to support decision-making processes in block cave design.

Aug 1, 2013

By P. Espinoza, I. Yutronic, J. Olivares

"Compañía Minera Doña Inés de Collahuasi&apos;s (CMDIC) Concentrator Plant is located at 4300 meters above sea level, approximately 2000 kms north from Santiago, Chile. The concentrator has three grinding lines. Each one, Lines 1 and 2, have one SAG mill at the primary grinding stage followed by a secondary grinding and classification stage consisting of one ball mill, two hydrocyclone batteries and a sump with two pumps. Line 3 has one SAG mill (Gearless Mill Drive), followed by a secondary grinding and classification consisting two ball mills, four hydrocyclone batteries and a sump with four pumps. Average throughput for these three lines is 150 Ktons per day. Peak values of 180 ktons per day have been achieved.Since late 2008 and throughout 2009 and 2010, Operations and Automation areas developed an optimization program for the grinding process, considering the implementation of Model Predictive Controllers (MPC) for the three SAG mills, and later, extending MPC to the secondary grinding and classification stages. The implementation of a MPC solution was based on stability considerations (high variability of the SAG process, high interaction between sump level control and classification efficiency, etc), requirements for an automated and consistent operation, the need to maximize throughput under timevarying scenarios and, in the long term, optimization of grinding process considering Specific Energy Consumption, among other parameters. MPC solutions had proven control robustness and strong optimizing capabilities in terms of helping in discovering new operational points and so achieving all of the aforementioned goals.On the first stage of the program, MPC implementation considered identification of dynamical process control models based mainly on historical data, in order to know the impact of a group of Manipulated Variables over Controlled Variables. Also, the impact of measured Disturbance Variables was modeled in order to take into account the high impact and interaction of these variables over the processes. At the end of this stage, 4 controllers were successfully operating for the 3 SAG mills and the secondary grinding and classification process. In response to challenges made by the working group, the second stage of the development of control applications in the grinding process required to achieve a design and implementation of a solution capable of governing the milling process, in addition to stabilization and optimization of both processes (SAG milling and secondary grinding), to integrate from a point of view of process control and automation. Therefore, in the second stage, MPC performance was continuously monitored and optimized through advanced tuning, inclusion of more process models (with proper step tests), redefinition of specific optimization criteria and adjustment to new process needs.Finally, successful MPC implementation at Collahuasi has proved that fast development and deployment of process control solutions is possible and allows immediate process stability, higher throughput, specific energy consumption reduction, and successful product size control, among many others."

Jan 1, 2012

By J. A. E. Bell

"Inco has shifted part of its traditional extractive research effort at its J..Roy Gordon Research Laboratory in Mississauga toward new, value-added products. Most of the effort is being expended on new Ni powders for batteries and new Ni-coated products which will be produced in Inco&apos;s carbonyl nickel refineries. New products under development are Ni-coated powders, Ni-coated fibres and ultra-fine Ni powders. Other new products for advanced material applications include controlled expansion composites, NiGOLD strip, Thermal barrier coatings on Inco alloys, and plasma coatings.Research and Corporate Strategic ThrustsThe newest strategic thru st of the company is to increase the value of the primary metal products. These new products are to be in addition to any new alloys or downstream alloy fabrication.Historically Inco has had the highest research expenditures in the extractive metallurgy field in North America. This has resulted in many new processes including flash smelting , matte separation, fluid bed roasting, pressure carbonyl refining and cobalt electrowinning which have contributed to establishing Inco as the lowest cost nickel producer. The recession of the 1980s caused a restructuring of Inco research and corporate goals .As shown in Figure I, from 1975-1982, i.e. pre-recession, Inco spent around 33 million dollars per year on research. Some 48070 of that sum was spent in Corporate Laboratories in Sterling Forest in the United States and in the United Kingdom , mostly on alloy associated research. A further 10% was spent on alloy research at the alloy production facilities in Hunting- ton, West Virginia, and Birmingham and Hereford in the United Kingdom. On extractive research, 13% was spent by process technology groups at the plant s, 20% at the pilot plants in Port Colborne and 9% at the J. Roy Gordon Research laboratory.The relative research expenditures (Fig. 2) have decreased so that today they are at the same level as 1975-80. The research as a per cent of sales has been cut from 2.2% to 1.2% of normalized sales. A normalized sale price of $4.00/lb Ni was used in the 1988 figures. This was partly achieved by closing Corporate Laboratories. Most of the essential alloy research was moved to producing plants and total alloy research is now 27%.A new thrust initiated in the 80s was mining research which now accounts for 20% of research expenditures. Over a period of the past five years this research has doubled the mining productivity in Inco. The pilot plant expenditures have been cut in half. One entire research station at Port Colborne was converted into a platinum group metals intermediate refinery. The decrease in pilot plant expenditures has been compensated by full scale plant trials devoted to the S02 abatement program at Sudbury."

Jan 1, 1990

By P. Erskine

"Hyperaccumulator plants can be utilized as ‘metal crops’ in agromining operations. This emerging technology produces ‘bio-ore’ by growing ‘metal crops’ on sub-economic ore materials, such as natural ultramafic soils. In a typical application plant material is periodically harvested, followed by air-drying and incineration to produce the bio-ore intermediate. The bio-ore is of very high grade (>15 wt. % nickel), and of high purity, and may be processed into a number of different products (including nickel metal, nickelbased catalysts and pure nickel salts). Current research efforts from our team focus on Mediterranean climate field trials (Albania, Spain) and tropical climate field trials (Malaysia, Indonesia). Field-scale demonstrations are required to provide evidence of real-life performance and of economic viability. If the trials are successful, agromining may in the near future support local livelihoods with income opportunities as an alternative type of agriculture: to farm nickel.INTRODUCTION AND BACKGROUNDHyperaccumulators, plants with the exceptional ability to uptake and concentrate trace elements in their shoots, can be utilized as ‘metal crops’ in agromining operations (Van der Ent et al., 2013a; Chaney et al., 1998; 2007; Baker et al., 2010). Agromining produces ‘bio-ore’ from harvested ‘metal crops’ grown on unconventional resources such as natural ultramafic soils (Van der Ent et al., 2015a). The bio-ore is a high purity and very high-grade (>15 wt. % nickel) product intermediate that may be turned into a range of different nickel products (nickel metal, nickel-based catalysts and pure nickel salts). The potential for agromining is greatest in tropical regions (Asia-Pacific Region: Indonesia, Malaysia, Philippines, Papua New Guinea and New Caledonia) where very large expanses of ultramafic soils exist (Van der Ent et al., 2013b). Agromine operations may replace existing marginal agriculture on poor ultramafic soils, for example oil palm plantations that require high fertilizer inputs to be profitable, or take place under a cocropping regime. The ability of ‘metal crops’ to achieve high nickel concentrations in biomass (and hence in harvested bio-ore) from soils with just <0.5 wt. % nickel, makes it possible to access non-conventional resources, not accessible by traditional mining techniques. For example, agromining could be implemented both on the large ‘halo’ around the mine leases, and as part of the progressive rehabilitation process after conventional resource extraction. As with all methods for resource extraction, agromining will be finite due to the diminishing concentrations of nickel in the zone accessible by plant roots (Chaney et al., 2014). Nevertheless, considering soil materials with 0.2 wt. % nickel and ‘metal crops’ with a yield of 100 kg nickel ha-1 the agromining venture may be sustainable for decades (Van der Ent et al., 2015a)."

Jan 1, 2016