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By Andrew Neale, Len Anderson, Rob Perry

"P.T Freeport Indonesia's new SAG mill commenced operation in February 1995. The reclaim system consists of four apron feeders and a series of three conveyors. The feeders are used in various combinations and relative speeds to blend SAG mill feed. It was recognized during design of the control system that effective automatic control would require dead-time and gain compensation. The successful implementation of the Dahlin algorithm to provide variable dead-time and gain compensation is described and compared to Pl control.RésumeL'opération du nouveau SAG de P.T. Freeport Indonésie a commencé en Février 1995. Le système de réclamation du minerai comprend 4 alimentateurs à tablier et une série de 3 convoyeurs. Les alimentateurs sont utilisés dans une combinaison variée et à des vitesses relatives différentes pour mélanger l'alimentation du SAG. II a été reconnu pendant la conception du système du contrôle que le contrôle automatique demandera une compensation pour le temps de délai et le gain. L'implantation de l'algorithme de Dahlin a été un succès et ii est comparé avec le contrôle Pl.ResumenP.T. Freeport Indonesia inició la operación de un molino semi autogeno (SAG) en febrero de 1995. El mineral es alimentado desde cuatro alimentadores tipo apron feeders y una serie de correas transportadoras. Los alimentadores se utilizan en variadas combinaciones y velocidades para mezclar apropiadamente el mineral que alimenta al molino SAG. Durante el diseño del sistema de control se detectó que el control automático requería de un tiempo muerto y compensaci6n de ganancia. En este documento se describe la implementación del algoritmo de Dahlin, que ha sido exitosa y ha permitido proveer tiempos muertos variables y compensaci6n de ganancia. Ademas se hace una comparaci6n con el Pl de control."

Jan 1, 1996

By Jacob Wood, Robert Matusewicz

The Outotec® Ausmelt Top Submerged Lance (TSL) Process has achieved widespread commercial acceptance over the past 40 years for the processing of copper, nickel, lead, tin, zinc and precious metal bearing feeds. The continual development and evolution of the technology during this period has resulted in a more intense smelting process with higher energy efficiency and lower operating costs. More recently, considerable effort has been invested in decarbonisation of the process through improvements such as increased oxygen usage, dry feed injection and use of low-carbon auxiliary fuels. This paper covers a number of these recent advances to the technology and highlights their impacts on energy consumption and CO2 emissions for a range of non-ferrous smelting flowsheets. INTRODUCTION Relentless demand for improvements in the profitability, safety and environmental performance of non-ferrous smelting plants has driven a continual development and evolution of processing technologies in this industry. In response to these demands, a number of bath smelting technologies have emerged over the past 40 years and have been implemented for the treatment of primary and secondary feedstocks. The Outotec® Ausmelt Top Submerged Lance (TSL) Process (henceforth Ausmelt Process) is one such bath smelting technology which has been widely adopted in a large number of greenfield projects and brownfield plant modernisations (Wood & Hughes, 2016). Emissions of CO2 from the production of non-ferrous metals represent approximately 4-5% of global totals (Rankin, 2012), comparable with the almost 10% attributed to iron and steel production. Despite significant improvements in the energy efficiency of modern smelting technologies (Coursol, Mackey, Kapusta & Cardona Valencia, 2015), primarily through increased oxygen usage, further innovation and development is still required to address their reliance on fossil fuels such as coke and coal. Transformation of the non-ferrous smelting industry to low-carbon technologies and process flowsheets is lagging behind the power and transport sectors, which have witnessed significant growth in clean/renewable energy technologies to reduce CO2 emissions and address the issue of anthropogenic climate change (IRENA, 2018; IEA, 2018).

Jan 1, 2019

By O. Guyot, A. Broussaud, M. Revalor, G. Legrand, D. Kok, E. Roux

"The paper describes the advanced grinding control system implemented at Newmont Ahafo gold mine, Ghana in 2009 and 2010. It shows how the different components of the Advanced Process Control (APC) system cooperate for increasing plant performance.The SABC circuit processes hard rock from different pits mixed with a variable proportion of oxides. The control strategy is designed to address specific disturbances, including start/stops of pebble crushers, unavailability of oxides, changes in rock size distribution resulting from segregation in the stockpile, and changes in ore type and hardness.The system consists of an expert system per se with extensive use of fuzzy logic, supplemented by an online dynamic model of the process used as a soft sensor, a VisioRock™ rock size measurement system over SAG mill feed before addition of recycled pebbles, and an AudioMill™ system to analyze the SAG mill sound and detect impacts on liners. All these components are embedded in the same OCS© software platform, which, among other benefits, makes it easy to support the entire system remotely. The paper details the physical constraints and the variables that are manipulated. It describes the control strategy and the contribution of each component of the system. It also describes the project implementation and the performance of the system. In addition to a strong involvement and support of management, the very good acceptance of the operators has been facilitated by the remote access system that made it possible to immediately modify the system each time it was needed during the baby sitting phase.The system made it possible to increase gold production significantly through increases in tonnage and recovery thanks to a better control of the grind size through a stabilization of the secondary ball milling circuit and an improvement of the operating conditions of the classifying cyclones."

Jan 1, 2012

By E. H. Hall

"Perhaps .I should start this talk on Arsine poisoning, by giving a brief background and history of the Bralorne Mill.The present Bralorne company started milling operations in 1932, with amalgamation and flotation, and operated continuously until 19 61 . In 19 59, it was decided to change over to cyanidation, and to use double filtration rather than counter current. The cyanide portion of this new plant was completed and went on stream in August 1961 utilizing the grinding circuit of the old flotation plant. The new grinding circuit was completed and started operating December 15, 1961.The grinding circuit consists of a 7 1 x 101 Hardinge Rod Mill, and a 7 1 x 101 Hardinge Ball Mill. In series with the rod mill is a 16"" x 24"" Denver Duplex Jig, and two of these jigs are also in parallel with the ball mill. The rod mill product is classified by a Kreb cyclone, the oversize going to the ball mill. The ball mill is in closed circuit with this same cyclone. The jig products are amalgamated daily, and the amalgam barrel tails are returned to the ball mill for regrind. Arsenious oxide (As203) is one of the reagents used in amalgamation. Approximately 39 ozs. per day is used.The cyanide circuit consists of a 481 x 121 thickener, three 24' x 24' agitators, two 141 x 161 drum filters and a Merrill Crowe precipitation unit.The Krebcone overflow is thickened to 60 - 62% solids. Hydrated lime, along with a settling agent is added to the thickener. Some of the overflow water from the thickener is returned for use in the jigs. The balance of this overflow water goes to waste. To the thickened pulp entering No. 1 agitator, barren solution is added, along with sodium cyanide and hydrated lime. Cyanide strength is held at 0. 9 - 1. 0 lbs/ ton solution and lime at 0. 8 lbs/ ton solution. The lime content usually drops to 0. 5 lbs/ ton in No. 3 agitator. The two drum filters are in series and are washed with barren solution. The discharge of No. 3 agitator is fed to No. 1 filter and the discharge of No. 1 filter is repulped in barren solution and feeds No. 2 filter. No. 2 filter discharges to tails, The filtrate is clarified, deaerated and the gold precipitated. One hundred to two hundred tons of barren solution are bled to waste each day. The ore contains 1 % Arsenopyrite and 1. 5% Pyrite."

Jan 1, 1964

By J. D. Templeton

A petroleum well that is not in use because it has ceased to produce oil/gas economically and dry holes are referred to as abandoned wells. Once a well is abandoned it is plugged with cement and can take anywhere from two days up to a week to cure depending on the number of plugs in the well. Since plugging a well costs capital, most wells are abandoned at the lowest possible cost and with the minimum obligations set by regulation agencies. These wells can present an enduring liability to petroleum companies. However, the depth and abundance of abandoned petroleum wells makes them an economically attractive source of geothermal energy. Geothermal energy harvested from an oil/gas well is a renewable energy source as long as it is extracted in a sustainable manner. The energy can be used to generate electricity, used directly for heating, incorporated into a water desalination process, or used by a heat pump for heating/cooling applications. In this paper we examine the possibility of extracting geothermal energy from abandoned oil/gas wells by studying the heat transfer in underground geothermal heat exchangers installed in these wells. The design configurations for the geothermal heat exchangers embedded inside a petroleum borehole can be a u-tube design or a shell and tube design (i.e. double pipe). Using in-situ gathered information from some representative petroleum boreholes, the effects of key parameters such as geothermal gradient, ground temperature values, and the flow inside of the tubes are evaluated. In order to provide a constant power production the inlet temperature can be adjusted to keep the difference between outlet and inlet temperatures equivalent. Higher differences between inflow and outflow temperatures will result in a less sustainable load put on the in-situ geothermal energy. Adding insulation to certain sections of the designs can minimize the transfer of thermal energy from the fluid to the surroundings. Insulation can also limit the thermal loss between the inflow and outflow tubes in the shell and tube design. Minimizing the energy loss of these designs leads to an increase in the outlet temperature, meaning a higher grade of heat. The sustainability of long term geothermal heat extraction is a balance between the rate at which geothermal energy is extracted and the rate at which the ground formation can replace this heat. Higher outlet temperatures can lead to a larger amount of power generation, and a higher coefficient of performance for heat pump applications.

Aug 1, 2013

By E. G. C Charlesworth, M. J. Robertson

"Gold mineralization in the Main Reef Complex at Sheba Gold Mine is hosted within shear zones in a sequence of shales and greywackes of the Fig Tree Group. Competency contrasts between these rocks and a steeply dipping chert unit provided a locus for the development of right lateral steeply dipping shear zones. A set of flatter dipping shear fractures crosscut these zones with their intersections defining a system of en échelon ore shoots. Mineralization is both of vein type and disseminated strata-controlled replacement type. The latter occurs mainly within the wall rocks to the shear zones. Arsenopyrite and pyrite are the major sulfides, with arsenopyrite being confirmed largely to greywacke beds and pyrite to the more iron-rich shale units. Gold was introduced, together with minor sphalerite, pyrrhotite, chalcopyrite and early arsenopyrite, toward the end stages of a first pyrite phase. The mineralizing fluids were introduced along dilated cleavage planes and, emanating from these, produced a series of replacement fronts. Mineralization was syntectonic to late tectonic and occurred under conditions of progressive deformation, resulting in sulfides which are concentrated in the hinges of shear-related folds and semimassive sulfides within tension fractures which are folded, attenuated and fractured. An analysis of the structures shows a sequence of events in which an initial ductile simple-shear event with a steep sense of movement was succeeded by a more brittle strike-slip event. Mineralization is temporally related to the latter. This evolution in structural style was probably related to reactivation of the Sheba Fault as a transcurrent fault during emplacement of large granitoid bodies."

Jan 1, 1994

By R. Sharma, I. Miskovic, S. Miskovic

Solvent extraction (SX) is a widely used method for the extraction of copper from pregnant leach solutions or other ionic solutions in hydrometallurgical applications. About 17% of total refined copper production worldwide is produced using SX followed by electrowinning (EW). Conventional SX units, such as mixer-settlers, consume a large amount of extractant, offer poor process control and yield low extraction efficiency. Extraction efficiency is poor due to low specific interface area, crud formation, emulsification, and slow phase separation stage. Micro solvent extraction (micro-SX) is a rapidly developing technology that provides an opportunity for improved process control, higher extraction efficiency, lower residence time and, as such, has a potential to reduce the environmental footprint of SX operation. In this research, extraction of copper ions in an acetate buffer media is investigated using D2EHPA with kerosene as a diluent. Chemical factors such as extractant concentration and ionic concentration of aqueous phase are varied to determine the optimum conditions. Subsequently, the extraction efficiency and reaction time of new micro-SX reactors are compared against previous bulk SX studies. Extraction in the microfluidic reactor was observed to be faster with a residence time more than an order of magnitude lesser than bulk SX. This study forms a prelude for developing new modular micro-SX units which will enable continuous, effective, and safe operation on a large scale. In future studies, numerical and analytical models validated through this work will help eliminate rigorous experimentation and enable faster design and fabrication of micro-SX units. INTRODUCTION Conventional industrial solvent extraction (SX) units such as mixer-settlers, extraction columns present various disadvantages. Low specific interfacial area, crud formation, and emulsification in industrial reactors lead to long extraction times and severely hamper extraction efficiency (Priest et al., 2011; Chen et al., 2017). Microfluidic extraction considerably improves upon the current extraction practices by (1) creating a large surface to volume ratio that enables rapid mass transfer across the interface, (2) better process control, and (3) easier separation of phases (Kashid, Renken, & Kiwi-Minsker, 2011; Ciceri, Perera, & Stevens, 2014). These factors establish micro-SX as a crucial mode for process intensification. Although, a micro SX unit has significantly low processing volume than conventional extraction methods, their throughput can be increased via massive parallelisation (‘numbering up’), which is an area of active research (Darekar, Singh, Mukhopadhyay, & Shenoy, 2016a; Kriel, Woollam, Gordon, Grant, & Priest, 2016).

Jan 1, 2019

By Alexander Levran

The main thrust of this presentation is to illustrate the growing contribution of solar energy to the mining industry by replacing existing mechanical and pneumatic tools and equipment with more effective and efficient electrical tools due to the increased efficiency and lower cost of solar energy generation. LCOE (levelized cost of energy/electricity) for solar energy is becoming highly competitive with other conventional sources of energy. The distributed architecture of solar power plants allows the supply of energy to highly remote and abandoned geographies. We will focus on the trend for this contribution to the mining industry over the next three years. Water desalination and delivery to the mining area will be topic of discussion regarding the contribution of solar systems in Latin America (LATAM) and other geographies. INTRODUCTION Increasing price pressure on the mining industry has generated a strong interest for mining companies to investigate energy costs and consumption. Currently, energy costs in mining economics contribute15% to 25% of the total commodity cost. In the past, most mines were operated by diesel gensets. Very high logistics costs for diesel supply, high and fluctuating diesel prices and theft led to very high energy prices. To overcome this problem, the mining industry has turned to solar energy, employing two main solutions: (i) adding a small scale (single level MWp energy supply) solar to an existing diesel system to create a hybrid solution operating as a microgrid environment; and (ii) replacing the existing diesel system with a large scale grid connected photovoltaic system financed by utilities or private IPP owners that signed a long-term power purchase agreement (“PPA”). In both applications, the near-term trend is to add storage and weather controls to optimize the electrical performance of the system. Use of concentrated solar solution (CSP) in the past was employed in a few remote sites to provide electrical energy to the mines. Currently this technology is not competitive with the photovoltaic (PV) systems. In some installations we have seen usage of a small hybrid system that is designed as relocatable, containerized and delivers scalable electrical power. In addition to the evolution of electrical energy sources, we see the trend of reduction of energy consumption in the mines by optimizing energy efficiency and generating load shifting solutions. Many studies of electrical energy in the mining industry support the growth of this trend in the mining industry.

Jan 1, 2019

By T. C. Chudy, C. C. Hamilton, G. R. Lane

The selection and provision of mineralogical services to assist in developing a GeoMetallurgical program for a Cu Porphyry Mine in South America is covered in this paper and presentation. The objectives of the study were to provide a rapid mineralogical scoping protocol to support and confirm results of sequential Cu leach data and to provide detailed mineralogical analysis of four composites and products on which locked cycle test-work were performed. Results showed good reconciliation was achieved between measured chalcopyrite and secondary Cu mineral contents with leach data using a linear intercept mineralogical procedure by TIMA (Tescan Integrated Mineral Analyser), as well as a generally positive correlation between pyrite and Cu-sulphide grain size. Size-by-size analysis of Composite feeds and test products was used to outline comparisons between feeds on a mineralogical basis and provide liberation and association data with which to diagnose and better explain performance. Identifying key mineralogical drivers to Cu recovery and concentrate grades using these results, a greater understanding of the role of chalcocite and pyrite on metallurgical performance was established. INTRODUCTION As a global supply of copper, porphyry copper deposits are typically very large, often in excess of a billion tons of ore (Keeney, Walters, & Kojovic, 2011; McMillan & Panteleyev, 1980) with grades typically between 0.2 and 1.2% Cu, and up to 2% Cu in places. They are also a source of gold. The complexity of copper porphyries arises from their classically zoned nature, the type of host rock, as well as whether or not they have been overprinted by oxidation, weathering and supergene enrichment (Sillitoe, 2011). Zones of alteration and other geological attributes give rise to varied mineralogy, ore mineral textures and material properties, all of which result in different processing characteristics. Due to the interdependencies of geological and processing characteristics, it is vital to study links between the geology at the macro-scale and mineralogical features down to millimeter and micrometer scale (mineral content and texture) and their influences on mineral processing behaviour (Cropp, Goodall, & Bradshaw, 2013). The growing field of GeoMetallurgy aims to provide an Orebody Knowledge framework by which to gather such information as it pertains to: (1) Understanding and predicting likely flotation response; (2) Guiding relevant test work to investigate and validate links, interdependencies, theories and data interpretation/evaluation; (3) Populate a block model with relevant information and; (4) Implement and refine the model for ongoing optimization and control. Specific questions addressed by using mineralogy often include flotation diagnostics as aids to interpretation, causes of dilution and losses to tailings, as well as guidance to sampling for test work (Cropp et al., 2013).

Jan 1, 2019

By Laxman M. Amaratunga, David N. Yaschyshyn

"In the mining industry several types of fills are used for backfill operations. They are broadly categorized into three basic types: rock fill, hydraulic/slurry fill, and paste fill. Recently, paste fill has become the popular alternative for backfilling because of its many environmental and operational advantages including its ability to use total tailings, both fine and coarse fractions.Amaratunga and Annor (1989) first presented the concept of agglomeration of fine tailings, which are otherwise excluded from backfilling, for utilization in backfill as a readily available supplementary fill material. Previous experimental work investigated the use of agglomerated fine gold mill tailings in low density slurry fills (60 -70% pulp density), consisting of varying agglomerate and tailings compositions.This paper presents the ongoing research into utilizing these agglomerates in a high density paste fill (70 -85% pulp density). Various binder dosages, curing period and combinations of pellet to tailings ratios were studied. Slump of the raw fill and compressive strength of the cured fill were examined. Results from the above study indicate that agglomerated tailings paste fill (ATPF) has superior strength characteristics when compared to total tailings paste fills.IntroductionThe Canadian mining industry produces more than 500 million tonnes of tailings and waste rock annually. These tailings consist of a coarse fraction and a fine fraction and can be either reactive (acid generating) or non-reactive. Most mining operations utilize the coarse fraction for backfilling while the fines have been traditionally disposed of in surface tailings ponds. Recently, paste fill has been developed which uses a fraction of the fine tailings (10 - 30% w/w of >45 µ,m) in the fill. This development allows a mining company to use a portion of the fines for backfill, but still requires surface impoundment of the unused portion of fine tailings. Reactive tailings have the added problem of generating acid after being oxidized in the presence of air and water. Treating acid mine drainage (AMD) seepage from these ponds has proven to be a costly environmental problem and long term liabilty.One practical method to reduce the environmental impacts associated with the tailings pond and to modify the physical and chemical properties of the fine mill tailings so that they can be utilized and disposed of in an environmentally safe manner through surface or underground applications is through agglomeration using the Cold-Bond Tailings Agglomeration technology (CBTA). Amaratunga and Annor (1989) first presented the concept of agglomeration of fine tailings for use as coarse aggregates for underground backfill, gravel or surface covering material. CBT A reduces the necessity of impounding tailings on surface while providing a suitable coarse aggregate resource material for practical use. Although the CBT A technology has been mainly limited to pelletization agglomeration alone, other forms of agglomeration like pressure agglomeration which include the techniques of briquetting and extrusion have also proven successful (Amaratunga, 1995)."

Jan 1, 1996

By S. E. El-Alfy, John Starkey

"The Iron Ore Company of Canada treats up to 125,000 t/d of iron ore at its Carol Concentrator in Labrador City, NF.A complete water balance study in the concentrator was done to show how to reduce the consumption of fresh water, to improve metallurgy in the magnetic separation (mag) plant, and to help reduce maintenance costs for the reclaim water and tailing pumping systems.The study showed that more than 60% of the water used in the mill was internally recycled dirty water from dewatering cyclone overflows, and that water reduction was closely tied to the effective use of these internal reclaim pumping systems. In addition, new operating parameters were set for the 4 -15,000 USGPM reclaim water pumps which should save over $100,000 per year in maintenance costs.IntroductionThe study described in this paper was set up for the purpose of lowering water usage in the concentrator, and to reduce tailing and process water pumping costs (operating and maintenance). Another benefit was also expected, namely, to improve magnetic separation efficiency through better water and density management of magnetic separator feed streams.The procedures selected to perform this work were: ( 1) to prepare accurate, as built flowsheets of the entire iron recovery process, including fresh and process water systems, (2) to calculate normal material balances in all circuits at full production, (3) to use the computer spreadsheet model so generated, to evaluate specific ways to reduce water consumption and (4) to re-evaluate metallurgical performance and operating parameters in the light of this information.Overall control of the 28 spiral iron recovery lines was found to be excellent because automated start up and shut down of the individual lines had been incorporated in the design from the start. Because production rates can vary due to sales requirements, mine and mill maintenance delays and hardness variance, the mag plant (which treats spiral tailings) must deal with a wide variety of operating conditions. Good water management therefore is a key item in achieving high recovery of the fine free magnetite which is too fine for recovery in the spirals.In order to make the study as accurate as possible, process data for major flows was reviewed by metallurgical staff and then by mill supervisors. The operating picture of the concentrator so generated was judged to be a good estimate of the mill at full production. By changing key inputs, the variance of process water demand and tailing slurry produced for several other operating conditions were studied.The usefulness of this exercise has prompted this present presentation to be made. Older plants by nature have undergone many revisions as operators have worked through ore changes, market changes, labour cost increases and the availability of new equipment, to improve their operation. Sometimes therefore it is necessary to stop and reassess what is happening in a plant in order to continue the never ending work of process optimization. The techniques used here were ideally suited to the gravity/magnetic concentration flowsheets used, but there is no reason why similar studies could not be done in other process plants.The authors of this paper hope that this experience at IOC will be helpful to others who face similar challenges in the pursuit of excellence in their processing plants."

Jan 1, 1995

By Damian Connelly

The option to leach iron from low grade copper concentrates could reduce transport costs and significantly improve the saleability of the concentrate. In addition, it reduces smelting costs and can remove undesirable impurities. The trend to produce cleaner and higher grade copper concentrates is very appealing to the smelters. A recent copper project came on line and there were three serious deficiencies which were not identified in the feasibility study. This came as a surprise and resulted in a number of projects to investigate ways of rectifying the problems in the shortest timeframe possible and using minimal new capital. The first serious issued not identified when commissioning was the lower than expected copper recovery, which was 70% compared to a predicted recovery of 80% to 87% in the feasibility study. This required investigating new flotation reagents and incorporating controlled potential sulphidisation. The second serious issue overlooked is the presence of significant amounts of talc minerals (insol) in the ore which are naturally floatable. This leads to lower concentrate grade and attempts to remove the talc resulted in lower copper recovery. Smelters have limits on the amount of talc in concentrate because it produces viscous slags and difficulties for the smelter. It also incurs higher transport costs. The third serious issue not identified in the feasibility study was the high arsenic reporting to the concentrate. Arsenic is a penalty element for smelters and at levels exceeding 0.3% the smelters may reject the concentrate. Arsenic is a problem environmentally for copper smelters and penalties apply when it is high. This paper describes the options for penalty element removal and suitability to incorporate into the modified flowsheet. It also highlights where the geometallurgy was not properly addressed and how these surprises could have been avoided. INTRODUCTION From 1970 to 1985, a copper company undertook metallurgical testwork on the Peruvian project copper ore. The ores were mined underground and treated over many years previously. Various other groups undertook studies and work until Aker Kvaerner undertook a feasibility study in 2006. Three ore composites were tested and grind-recovery relationships were developed. The feasibility study testwork identified that talc was a possible concern and will have an impact on copper concentrate grade. This was to be managed by blending. The current operating reality is that there is so much talc in the ore body it is not possible to blend it out. There was far more emphasis on molybdenum recovery. No mention was made of arsenic issues other than blending. There was a Concentrate Leach Study and a Molybdenum Hydrometallurgical Plant Study but these options did not proceed to a final flowsheet. A Scoping Study in 2006 looked at a Heap Leach project. The heap leaching gave very poor copper recovery on the sulphide ore. Heap Leaching was evaluated and tested for the high talc and high arsenic ore but recoveries were poor. The Aker Kvaerner feasibility study opted for a simple primary crush SAG/ball grind followed by flotation and sale of a concentrate. This is the current flowsheet utilised and the scale of the project is very large based on processing 117,200 tonnes of ore per day. A feasibility study does not guarantee success and the deviations from the feasibility study suggest there is a lot of technical work required to improve the current metallurgical performance.

Jan 1, 2019

Ni and samaria-doped ceria (Ni-SDC) cermet are desirable materials for making a solid oxide fuel cell anode working at intermediate temperature (800°C or lower) due to its good ionic/electronic conductivity at the reduced operating temperature. An advantage of the use of a Ni-SDC anode is the enlargement of the effective reaction zone in the interfacial area between the anode and the electrolyte. As a result of the mixed ionic/electronic conductivity of the anode, the electrode reaction rate is enhanced. Investigations into the properties of the ceria-based anode reveal that the performance of the anode is closely related to its microstructures, which are dependent on the composition and fabrication conditions of the ceramics. In this paper, a method for quantitative characterization and correlation of Ni-SDC cermet microstructures to the performance of an anode is investigated. A series of Ni-SDC anode samples are prepared by screen-printing (SP) and spin-coating (SC) techniques, and different microstructures of the samples are obtained through controlling the SP or SC conditions and the sintering temperature. Individual phases (Ni, SDC and pore) in the tri-phase boundary (TPB) microstructures in the electrode are differentiated and segmented from each other in optical and SEM images based on the differences in refractivity and atomic contrasts among the phases. The geometric properties of the segmented phases are characterized using image analysis methods, so that a quantitative skeleton illustrating the TPB microstructures of the Ni-SDC electrode is obtained. Statistical analysis of the distribution of the three phases in the TPB microstructures is performed and the relationship of the microstructures with the optimum electrode is interpreted in terms of the microstructure analysis data.

Jan 1, 2005

By Christopher R. Davis

Resource modelling is the basis for any economic appraisal of a mining project and includes a number of steps from data acquisition and validation to resource reporting and classification. At each step in the modelling process it is necessary to define the specific objectives, the methodology proposed to achieve those objectives and to establish a set of checks and validation tools to assess the effectiveness of the proposed methodology. The data used has to be both reliable and relevant for the purpose of the modelling. QA/QC for database validation requires verification of sample location; including correct collar location, topographic modelling and down-hole surveying. Validation is also performed on assaying information such as checks for sampling bias and accuracy of core splits, proper use of pulp duplicates and standards, handling of calculated values, explicit/implicit absent data and data entry errors. The second step involves identifying the mineralized domains resulting from the mineralization process. This step requires a geometric interpretation of all relevant geological features that have influenced the spatial distribution of the mineralization including dykes, faults and alteration. In addition, a geologic reference system rather than a standard Cartesian system is established to measure distances between samples. Once the mineralized domains are established and the database has been validated, statistical and spatial description of the mineralized samples are examined in order to establish the histogram and the variogram of all variables of interest and their correlation. Results are also used to calibrate the global resource estimate within the mineralized domain and to establish the search strategy for grade interpolation. The mineral envelope is filled with polyhedral cells representing mining blocks. The average grade of these blocks is estimated using kriging techniques in order to eliminate global bias and to minimize the local error in the block grade estimate. Visual checks are made to validate the spatial trends in grade distribution while global bias and smoothing effect are assessed from the declustered sample statistics. When required, smoothing correction may be applied using ?change of support? techniques. Simulations are used to perform sensitivity analysis and the simulation models are used if severe smoothing of spatial variability makes the interpolated model inappropriate for mine planning. The keys to successful resource estimation are an integrated team of geologists, geophysicists, geostatisticians and mine engineers using the best available technology and reviews by independent auditors. The resource team is also responsible for performing a QA/QC program in order to measure the effectiveness of the methods in meeting the specific objectives at each step of the modelling process.

May 1, 2001

By Z. Xie

Gas diffusion electrodes (GDEs) for high temperature polymer electrolyte fuel cells (PEFCs) have been prepared using ZrP-modified Nafion ionomer. The performances of these GDEs as cathodes and anodes were evaluated and compared with conventional Nafion ionomer GDEs using Nafion 112 as the PEM under elevated temperature and reduced relative humidity. ZrP-modified Nafion ionomer was incorporated into the catalyst layer to enhance water retention and/or increase the protonic conductivity. The fabrication of ZrP-GDEs was implemented by the ion exchange of Nafion solution with zirconyl chloride, followed by the impregnation of GDEs in H3PO4 solution. The composite electrolytes in the catalyst layer possessed a ZrP content of 20% by weight. GDEs were characterized by TGA, SEM, XRD, and CV. Scanning electron microscopy (SEM) indicated evenly distributed elemental Zr and P within the catalyst layer. Cyclic voltammetry shows a reduced electrochemically-active surface area when the ZrP is incorporated into GDEs. Polarization and impedance analyses indicate the cathode with ZrP-modified Nafion ionomer did not show an enhanced or decreased performance, compared with conventional Nafion ionomer GDEs under elevated temperature and decreased humidity. However, when GDEs with ZrP¬Nafion ionomer were used as the anode, an improved performance was observed. Preliminary experiments show that the anode is more sensitive to dehydration than the cathode under elevated temperature and/or low humidity. Improving the anode water retention capability will help to reduce losses at the anode. However, the incorporation of hygroscopic agents in the cathode catalyst layer is not essential because it is self-humidified by electrochemically-produced water even under elevated operating temperature and lower humidity.

Jan 1, 2005

By G. W. H. Norman

Introduction During the past few years emphasis has been placed on the importance of major faults, fault zones, and associated structures in localizing ore deposition. This interest has developed partly through experience in mining camps in other parts of the world, and partly by the discovery of such structures associated with ore deposits in the Abitibi region, northwestern Quebec. The recognition of these structures has been brought about chiefly by detailed mapping that has been gradually extended along most of the southern gold belt of Quebec and across adjacent parts of Ontario. In the Chibougamau region of northern Abitibi, progress has been made in the mapping of major faults by using aeroplane photographs with a scale of about 1,000 feet to one inch for field mapping. These studies have shown that many large faults occur in the Abitibi region; also that there is a relationship, direct or indirect, between faults and mineral deposition. The faults are, however, of more than one type and age, and this should be clearly appreciated in any planned programme for prospecting this great region adequately. The information regarding faults and structure is so widely scattered that it is difficult to obtain a clear picture of the essential features, either for planned prospecting or for analysis and development of ideas for further study. This paper, as originally conceived, was intended only to discuss and compare the regional structure of southern Abitibi county-where detail mapping and diamond drilling have been carried on intensively for the past ten years-and of the less well-known Chibougamau district, in order that the experience gained in the one area could be usefully applied in the other. However, a comparison of the two districts leads naturally to a consideration of their common relationship to certain major late-Precambrian structures that form an important northeast system across the southeastern part of the Canadian Shield.

Jan 1, 1946

By A. A. Nourbakhsha

During the last few years, low cement castable refractories found more application in industry, because of good low and high temperature mechanical properties.In the case of alumina based castables ,usually fume silica or reactive alumina are used as superfines in the matrix to form mullite at high temperature which depends to source of alumina . Mullite network formation improves the hot mechanical properties and thermal shock resistance of bodies. The mullitisation of andalusite (so called primary mullite) also produces a certain amount of rich silica glass, part of it being expelled out of the grain. This phenomenon restrict the application of andalusite low cement castable in pig iron ladle. In this research, firstly 8 pig iron ladles from ESC(esfahan Steel company) with different conditions selected and monitored during usage to determine of main factors affecting on corrosion of refractories. Results showed that tap to tap time and undischarged pig iron are the main ones.Therefore monolithic refractories such as castables which increase thermal shock resistance could be substituted. Regarding to ternary phase diagram of Al2O3-SiO2-CaO ,fumed silica was selected as fine particle.After optimization of particle size distribution by using of modified andreasen equation and rehological properties ,different mixture were composed with andalusite ,fume silica, reactive alumina,secar 71 ,SiC and kyanite.All aggregates system components were mixed dry for 1 min in hobart mixer .After water addition wet mixing was continued until getting flowing about 4 min.The ready mixes were poured in the moulds of of 25mm X 25mm X 140mm with vibrations .After demoulding the samples were cured for 24 hrs and dried at 110OC/18hrsand then fired at 1100-1500OC. Physical ,mechanical ,phase analysis(XRD) and microstructure studies (SEM)were carried out according to ASTM standards. Results showed that by addition of kyanite ,more mullite phase formed through the matrix but,because of expansion occured during this phase transformoration, some cracks formed and mechanical properties decreased.By SiC addition the silica film on the surface of SiC lead to improvement of mechanical properties at high temperature

Oct 1, 2011

By G. Vibert Douglas

IN a previous paper by Douglas and Gillies (1), a simple field test was given for gold. Later work by the present writers, and by others who have kindly written to us, shows that the test is not as satisfactory as it was hoped. In many cases, the gold mirror is completely obscured by a dense precipitate which, up to the present, it has been found impossible to remove. After many attempts, we have temporarily given up the iodine test and have been working on a simple pyridine reaction based on the work of Putnam, Roberts, and Selchow (2): The ore to be tested is ground to a fine flour. Place half a gram of this powdered ore (an amount that will cover a Canadian ten-cent piece, not heaped) in a six-inch test-tube and add 5 c.c. of aqua regia. The aqua regia. should be made up on each occasion, using four parts of hydrochloric acid and one part of nitric acid, both concentrated. Shake the solution in the test-tube for about one minute and allow to settle until dear. The rime to settle may be 30 to 60 minutes. When dear, a drop of the solution is with-drawn with a fine, dean eye-dropper or capillary tube and placed on à small glass plate. A slide such as is used for mounting rock sections is suitable. Now place an equal sized drop of pyridine solution on the plate beside that of the aqua regia solution to be tested. The pyridine solution, containing nine volumes of hydrogen bromide to one volume of pyridine, can be made up in advance and carried in a glass-stoppered bottle. Join the two drops with a glass thread or fine, dean stick, and allow the merged drops to stand undisturbed. I t is well to place the slide on a piece of white paper so that the presence of the crystals which may form can be easily seen. The time for their formation will vary with the amount of gold present, but the slide should be allowed to stand for at least 30 minutes before negative results are reported. A hand lens of at !east ten magnifications is essential. If gold is present, minute red crystals will form around the margin of the drops as these drops slowly evaporate.

Jan 1, 1938

By W. T. Boyes, J. M. Harrison

"Squall Lake Gold Mines, Limited, owns a group of claims at Squall lake, northern Manitoba, about 75 miles east of Flin Flon and 3 to 4 miles north of the Nor-Acme mine at Snow lake. Gold mineralization was discovered in this locality several years prior to 1944, but it was only in that year that a comprehensive examination of the occurrences was begun. Development work to the end of 1946 has been encouraging and the deposits show interesting relations to rock structure.GEOLOGICAL SETTINGThe gold occurs in a sill of biotite-hornblende diorite at or near its contact with gneisses derived from arkosic sediments . In detail, the contact between the diorite and the arkose to the southeast of it is irregular, with interfingering of the two rock types. On the northwest side of the sill, between it and Squall lake, the rock is staurolite schist (see accompanying sketch-map).Several mineralized shear zones are known, of which those shown are the most important. Within the zones, the diorite is silicified and somewhat carbonatized. Fairly coarse arsenopyrite is the chief sulphide mineral in these zones but, locally, they contain also pyrite and pyrrhotite. Where quartz and chlorite are present, rather coarse gold is associated with them. In that part of the main zone where gold assays have been most consistent, the diorite is rich in albite."

Jan 1, 1954

By J. M. Cormie

"The orebodies of the East Malarcic mine are interesting from a geological standpoint because of the very small amount of vein quartz, the occurrence of much of the ore in soft greenstone schist, the presence of innumerable bodies of porphyry and diorite, and their obvious relationship to structural features .The property lies in the northeast part of Fourniere township in western Quebec, and within the important fault zone that geological mapping in recent years has shown is traceable wesrward through the Noranda and Kirkland Lake camps and eastward to beyond the Val d'Or region . The claims included in the property were staked by J.P. Norrie and L. K. Coffin in 1932, but gold was not discovered until i934. Production began late in 1938. The total amount of ore milled up to the end of 1945 was 2,837,091 tons from which gold having the value of $16,631,758 was recovered.The region is underlain by an alternating series of Early Precambrian sediments and volcanics cut by innumerable intrusives. These are more fully described in the works of Gunning, Ambrose, and Norman. The East Malanie property is underlain by three formations, which, named from north· to south, are : the Cadillac sediments, the Blake River volcanics, and the Kewagama sediments. In addition, there are countless intrusions of porphyry and diorite, particularly in the greenstone, occurring in all sizes up to those measurable in hundreds and even thousands of feet . These constitute a large proportion of the greenstone in places. In many parts of the ore zone there is more porphyry and diorite than schist. The greenstone itself is not homogeneous but includes peridotite, andesitic lava, and soft talcose schist.Orebodies have been found in both the Blake River volcanics and the Kewagama sediments. They are not known to occur in the Cadillac group. The ore occurs in a zone of pyrite mineralization in which almost no other metallic minerals and very little vein quartz are present. Generally, the grade varies directly with the amount of pyrite. The mineralization is later than all the rocks of the area. Diorite was a particularly favourable rock for mineralization and contains the highest grade of ore, but any rock in the fault zone may be ore-bearing, whereas the same type away from the zone is not."

Jan 1, 1954