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By Kamran Esmaieli

Open Pit Mining Through Underground Voids: Challenges Main Challenges/Hazards: Unexpected collapse of the pit floor and/or pit walls; The loss of equipment and/or people into underground voids; Loss of explosives from blast holes that broken through into the underground voids. How to address these challenges? An important step that must be integrated into mine planning and design is the accurate detection and delineation of underground voids. (i.e. probe drilling, Geophysical techniques, laser scanning,?); Leaving a pillar of adequate thickness between the working bench and the underground voids; Backfilling the underground voids from surface using vertical or inclined raises; Monitoring the working locations (pit floor and pit slopes) to detect movements.

May 1, 2013

By R. S. Fitzpatrick, P. Hegarty

"Gravity separation is a key technology in the recovery of a range of mineral ores including coal, tin, tungsten and tantalum. One of the problems associated with gravity concentration is that recovery is traditionally lower than for other separation techniques such as the flotation of base metals. Any improvements in recovery will, therefore, have a significant effect on the economics of operations of this type. More accurate, robust modelling of gravity separators is one means of achieving greater recoveries. Furthermore, better modelling can lead to more informed equipment selection and process design for potential economic deposits.Previous models for gravity separation equipment have typically relied on simple metal mass balances from samples of the separated products. However, with advances in technology there is a ready availability of analytical instruments for the automated quantitative determination of mineralogy. Instruments such as the Mineral Liberation Analyser and QEMSCAN make it possible to estimate particle size and density (related to mineral liberation and association) which are the primary factors in gravity separation. Models produced using quantitative mineralogy of feed streams will be potentially superior to those produced historically.This paper describes test work undertaken on a wet shaking table using a synthetic ore. The objective of this test work was to develop a model to predict the effect of changes in feed ore characteristics on shaking table performance by measuring quantitatively the mineralogy of the feed material. Particles were divided into size/density classes based on the results of QEMSCAN analyses.It was found that it was possible to use data from QEMSCAN to predict the recovery of size/density classes and to predict the effect of changes in the relative abundances of these classes on the separation. This model can potentially be used to adjust operating variables to maximise grade and recovery by monitoring the input feed. Care was needed when utilising QEMSCAN data to ensure representivity."

Jan 1, 2016

By J. R. Rodda

Stainless steels are widely used to handle concentrated sulphuric acid (H2SO4) solutions produced in the manufacturing process, The majority of corrosion rate data published for stainless steel in product acid conditions is typically complied in tables or in iso-corrosion diagrams (iso-corrosion rates plotted as a function of temperature and H2SO4 concentration), without any reported consideration as to the accuracy and relevance of the method used to obtain the data, Given the characteristic potential cycling associated with the active-passive corrosion behaviour exhibited by stainless steel in hot, concentrated H2SO4, weight loss is considered to be a more reliable method than electrochemical measurements to measure corrosion rates, Consequently, an effort is made herein to demonstrate the accuracy and relevance of weight loss to determine laboratory corrosion rates in product acid conditions.

Jan 1, 2010

By Mario E. Rossi

Geostatistical conditional simulations are becoming increasingly popular as tools that provide models of uncertainty at different stages of a mining project. They have been used as grade control tools in daily operations, to assess the uncertainty of minable reserves at the project?s feasibility stage, and to assess mineralization potential in certain settings. Other applications include assessment of recoverable reserves, resource and reserve classification, and drill hole spacing optimization studies. All large-scale applications of conditional simulations intend to benefit from the ?correct? (or ?accurate?) characterization of uncertainty, adequately describing the variability observed from the data and its impact on the final objective. The objective of this paper is to compare the ad-hoc ?uncertainty model? provided by a traditional classification scheme with the probability intervals resulting from a conditional simulation model. It is argued that the uncertainty model derived from the use of classification schemes is of little practical use to the mining company or operation. The paper emphasizes differences in uncertainty based on the size of the mining unit considered and the extraction method, and local geological differences. Most importantly, it discusses confidence intervals for these mining units, as compared to the relative precision envisioned by a typical classification scheme. The application described in this paper is taken from a medium-sized Cu open and underground mine in northern Chile, and it demonstrates the impact of several variables on the resulting conditional simulation models. It also compares a more ?traditional? resource classification scheme applied, and the wide difference in local confidence intervals for each mining unit.

May 1, 2001

By K. Scalise, M. Teixeira, K. C. Kocsis, J. Parrish

"This paper aims to introduce and discuss a newly developed algorithm, which can be uploaded on smart watches and other mobile devices to inform and protect underground personnel working in hot and humid environments. Using this algorithm, the metabolic rate (M) of a worker can be determined in “real-time”. In addition, besides warning signs and alarm signals emitted by the smart watch, the algorithm has the ability to determine work-rest time periods based on the worker’s physiological parameters (e.g. age, height, weight, gender) and the climatic parameters (Td, Tw, flow velocity). This paper also investigates whether a re-designed auxiliary ventilation system can provide appropriate work conditions in the production stopes. INTRODUCTION An underground environment with high temperature and humidity conditions generated from various heat and moisture sources can significantly affect the thermoregulation processes of the human body. This can lead to heat-related illnesses such as heat rash, heat cramps, heat exhaustion, and heat stroke. Heat stroke is a serious illness that carries a high risk of fatality, if the underground workers are not immediately treated and the climatic conditions are not corrected (McPherson, 1993). There are two methods used to control the climatic conditions in underground mines: (1) Using ventilation and cooling systems to reduce the thermal stress on the mine workers, and (2) Reducing the heat transferred to the ventilating air from major sources such as: strata, auto-compression, mining equipment, groundwater, blasting, etc. The heat from strata and auto-compression is transferred to the ventilating air at varying degrees, based on factors such as the virgin rock temperature (VRT), the thermal properties of the surrounding rock formations, and the surface air temperature. However, the use of mining equipment and blasting processes can be managed to reduce their contribution to the mines’ overall heat load. To provide appropriate work conditions, selecting the most suitable cooling method depends on the level of heat to be removed, the location of the problem areas, and economic constraints (Brake & Bates, 2002; McPherson, 1993). Because every mine is unique, it is essential to monitor the climatic conditions to select the most appropriate method of cooling. A majority of the time, simply adjusting or re-designing the auxiliary ventilation system can efficiently remove the heat and dilute the contaminants. This may involve moving the ducting system, changing the size of the duct, changing the size of the auxiliary fan, or switching from forcing to an exhausting fan arrangement. If the required air volumes are higher than the existing primary ventilation system can supply, other measures may need to be taken to increase the airflow delivery. This may include regulating the flow in some areas of the mine to increase airflow to the desired area, adding booster fans, or replacing the primary fans."

Jan 1, 2019

By K. Mayhew, K. Murray

The CESL Copper Process was developed as a hydrometallurgical alternative to smelting and refining for copper sulphide concentrates. As the majority of such concentrates contain appreciable gold and silver values, effective recovery of these metals is essential to the economics of this technology. Unfortunately, conventional cyanide processing cannot be used on the residue from the copper process, clue to its unusual characteristics when compared with naturally occurring gold and silver ores. An abundance of elemental sulphur, as well as the presence of cyanide soluble copper, leads to the formation of both thiocyanate and copper cyanide which is detrimental to process economics clue to excessive cyanide consumption. This, however, can be reduced to acceptable levels by process innovations such as pressure cyanidation and cyanide recovery. This paper outlines the successful application of the CESL Gold Process to various gold-bearing copper concentrates from bench to demonstration scale.

Jan 1, 2011

By S. Shairp

The first commercial application of the Dynatec two-stage zinc pressure leach process commenced operation at the metallurgical complex of Hudson Bay Mining and Smelting Co., Limited in Flin Flon, Manitoba in 1993. A key feature of the successful leaching operation is the presence of iron in solution, which facilitates oxygen transfer in the autoclaves and enables rapid oxidation of zinc sulphide minerals. At the same time, the solution from which zinc is recovered must be essentially iron free. Precipitation of iron from the solution starts in the autoclaves and is completed in subsequent atmospheric purification steps. Conveniently, these iron precipitation steps also remove nuisance elements such as arsenic and fluoride from the solution, while producing solids with good liquid-solid separation characteristics. This paper provides an overview of the current operations in Flin Flon, with particular reference to the behaviour and deportment of iron throughout the zinc plant.

Jan 1, 2006

By G. Ellie, D. F. Coates

"The mining of coal in thick and inclined seams is difficult, and there are only a few countries in the world that are really interested in the problems associated with their exploitation. In France a considerable amount of investment has been made in the mechanization of systems appropriate for such conditions.For seams inclined up to 50 degrees, of normal thickness, a new system was introduced in 1967. It consists of a longwall advancing to the rise with walking props, a shearer and hydraulic backfill. During the first two years the panel productivity was an average of 25 TN/MS, and they foresee the possibility of reaching an average of 30 TN/MS.A system for thick beds, both horizontal and inclined, is in p•rocess of being mechanized. In 1969 they started with a method consisting of a longwall in a footwall slice, or sub-level drifts with walking props, and caving and drawing up to a height of 20 m of coal. Recent experiments with a shearer have produced panel productivities between 9 and 11 TN /MS, and there is hope that with experience an average of 16 TN/MS will be obtained.In 1964 a method was initiated for beds inclined from 60 to 90 degrees, based on a mechanical cutter designed by the collieries in Lorraine. An overhand slice of 4.2 m in height is taken with a drum cutter mounted on the end of a boom. Hydraulic fill is placed at the completion of a slice. The face productivity, which has been from 13-20 TN /MS, can be perhaps increased to an average of 22 TN/MS. For those companies with important reserves in inclined and thick seams, these French systems may be worth the trouble of examining in some detail."

Jan 1, 1972

By L. B. Smith

Introduction The property of Wright-Hargreaves Mines, Limited is situated at the northeastern end of Kirkland lake, Ontario. It lies immediately east of the Lake Shore mine, and about in the centre of the Kirkland Lake camp. The district is served by the Swastika-Rouyn branch of the Temiskaming & Northern Ontario railway, the town of Kirkland Lake being four miles northeast of Swastika, which is 63 miles north of Cobalt and 392 miles north of Toronto. The mine is developed by three shafts and has more than twenty-eight miles of underground workings. There are two main producing veins, which lie in a porphyry boss 1,350 feet wide. Milling started in May, 1921, and to-date (January, 1933) the mine has produced gold to the value of $21,770,183. Present production is at the rate of 800 tons of ore per day. For a description of the geology, reference may be made to the report by E. W. Todd on the Kirkland Lake Gold Area (1). Structure and Ore Deposits The two major productive veins on the Wright-Hargreaves property are known as the North vein and the South vein. In the vicinity of the shafts these veins are approximately 550 feet apart. A third vein, known as the Inclined vein, has been worked from its junction with the South vein at the 2,000-foot level to its intersection with the North vein at the 800-foot level. This inclined vein has to date been found only in the area east of the eastern major north-south fault. The main North vein and South vein have a general vertical dip, although local rolls and post-ore faults throw certain sections out of line. From the surface to 3,900 feet, the North vein is displaced south less than 200 feet, and the South vein at the same horizon is almost perpendicularly beneath the surface showings.

Jan 1, 1934

By I. Yutronic

Compañía Minera Doña Inés de Collahuasi'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. E. Quinn, S. Wolstencrof, K. H. Soldenhoff

Ion Exchange (IX) using Strong Base Anion Exchange (SBA) resins is conventionally used to purify uranium from sulfuric acid pregnant leach solutions (PLS), particularly when uranium grades are relatively low. However, the presence of chloride in the PLS interferes negatively with conventional SBA based IX processes, by reducing the uranium loading on the resin. Chloride can be present in the PLS due to the use of saline bore or seawater in leaching, or chloride containing leach/IX reagents. ANSTO has been examining strategies to combat this problem for several years. ANSTO previously reported that a recently developed high-capacity SBA resin was capable of extracting uranium efficiently from saline leach liquors produced at Boss Resources’ Honeymoon in-situ recovery mine in South Australia. The resin was tested successfully in a 2017 Field Leach Trial carried out onsite in the existing wellfield area. While the piloting was extremely promising, the elution was more challenging, with eluate tenors limited to 2 g/L U3O8 using NaCl. This is due to the resin’s strong affinity for the uranium complex, which makes it difficult to elute efficiently by standard methods. In this paper, we describe a novel elution process developed for strong base resins that significantly improves upon conventional methods. The process, developed by ANSTO, is based upon conversion of the extracted uranyl sulfate species to uranyl carbonate, followed by elution with sodium chloride. The loading and elution process was demonstrated in continuous locked-cycle piloting conducted at ANSTO in 2019. The results were excellent, with eluate tenors of 8–9 g/L U3O8 achieved from Honeymoon PLS containing 50 mg/L U3O8, with very little transfer of impurities. In subsequent test work, the eluate was shown to be amenable to uranium peroxide precipitation. The results represent a huge step forward for the purification of uranium from saline leach liquors.

Jan 1, 2020

By J. P. Sairrlon, K. G. Wheeland, K. J. Mintz, B. J. Kiener

Iwenty-one ore dusts of varying mineralogy, sulphur content and reputed in-mine explosibility were selected from eight mines. After chemical and mineralogical analysis and sieving into two size fractions (+45/ -63 microns and -45 microns), these sarrples and seven pure minerals were tested for explosibility in oxygen in a Harbnann chamber. '!he maximum explosion pressures and maximum rates of pressure do not correlate with sulphur content alone; the types of minerals are also .important. Residues were analyzed to check the combustion reactions. Further comparisons between laboratory and mine results are planned to improve predictability.

May 11, 1988

By R. J. Enright

Eight mines in canada and three in Australia were visited between September 1987 and Februacy 1988. All mines had encountered problems with sulphide dust explosions and had introduced defence mechanisms to prevent or limit the extent of these explosions. Techniques included dispersion of limestone by various methods, use of water foggers, placing of water and inert stenuning in blast holes, and removal of settled dust to prevent secondary explosions. The techniques were evaluated and it was concluded that certain practices should be modified or discontinued as they were of little value in preventing dust explosions.

May 11, 1988

By J. Petit

"In July 2004 workers at the ABI Smelter in Bécancour (Quebec, Canada) went on strike. Contingency plan were ready to face the situation and plan was done to operate one potline out of three. Two lines were stopped; keeping in mind the restart on cold metal after the strike was over. The subject to be presented and shared includes a summary of work organization including training, schedule of restart, pot cleaning, pot preparation, crash start method, follow up after crash start, safety and environment. This major project has been a very big challenge. It shows that people involvement is a must and the key success. It is then true to say : ""People who believe in success will work and act in order to get there"" and ""People who don't believe in success will act in such a way to fail"" Then both will prove they were right to think their own way ! Sharing such experience can be interesting because it's a true story and people like true stories."

Jan 1, 2006

By J. Kabuya Mukendi, F. Brunet, J. Hadjigeorgiou, P. Matte, D. Leblanc, M. Grenon

La stabilité des bancs miniers est un élément important de la conception des fosses. Il n’y a pas de consensus sur les critères à utiliser pour évaluer la fiabilité des bancs miniers. Lors de la conception des fosses, la fiabilité des bancs miniers est analysée à l’aide d’un ou deux critères. L’objectif de cet article est d’analyser l’efficacité de la conception des bancs miniers de la fosse Tiriganiaq du projet minier Meliadine par rapport à des critères de fiabilité rencontrés dans la littérature et la pratique. L’article présente les similitudes et les différences entre les résultats obtenus pour les différents critères et propose une approche multicritère systématique pour évaluer la fiabilité des bancs miniers à l’aide de cartes de susceptibilité.

Jan 1, 2012

By H. A. Baxter

THE Eustis concentrator is located about nine miles southwest of Sherbrooke, on the main line of the Quebec Central railway. The concentrator is served by a side track parallel to the main line, and all supplies for both the mine and mill are delivered at the mill yards. The concentrator treats 250 tons of copper-iron ore daily by preferential flotation, whereby, first, the copper sulphides are floated and the iron pyrite is depressed, and, second, the iron pyrite is reactivated and floated, resulting in an iron pyrite concentrate and tailing, the latter going to waste. During 1934, 71,709 tons of ore were treated, averaging 3.712 per cent copper and 32.1 percent iron, with a small amount of silver and a trace of gold. The recovery was 93.77 percent of the copper, and 70.43 percent of the iron as iron concentrate. The copper concentrate produced averaged 25.623 percent copper and the iron-pyrite concentrate averaged 42.13 percent iron and 49.173 percent sulphur. HISTORY The first mill at the Eustis mine was built in 1904 (1). For a few years previous to this rime, some jigs had been operated to concentrate the fines from the mine, the coarse ore being hand-sorted. Both the fines and sorted lump ore were shipped to acid plants. The building of the mill was based on test work carried out by Professor Robert H. Richards at the Massachusetts Institute of Technology. The method of concentration consisted in careful classification in a Richards classifier, which made six different sized products. These products went to Wilfley tables, where a concentrate, middling, and tailing were made. The middling was returned to the rolls. Crushing was done by a jaw crusher and two sets of rolls, resulting in a screen size of about five-mesh.

Jan 1, 1936

By C. E. Baltzer, W. H. Harper

"This study has been developed to illustrate the trend and pattern of distribution applying to the quantities and costs of the fossil fuels used, and the related amounts of electricity generated, in the thermal segment of the Canadian electric utility industry.Traditionally, Canada's requirements for electric energy have been preponderantly met by hydraulic power. The need for augmented supplies of thermally generated power has mainly developed since the close of World War Il-in an era that has seen drastic decline in the production of indigenous coal and buoyant expansion in the supply of oil and natural gas. Although coal currently is faced with serious competition from the fluid fuels, the utility industry in particular promises an advancing market for coal, because of its abundant resource, at prices less likely to show sharp increase. For Canada as a whole, this trend has been progressive since 1959, and, unbelievable as it may now seem, the extent of the advance by 1975, particularly in Ontario and the Prairie Provinces, may well be as much, or even more, than the present rate of consumption for all purposes.The text and the tables and charts are arranged to focus attention on: the growth of thermal genera ting capability; the historical trend in consumption of fuel and the related production of electricity; patterns of distribution by types of fuel and thermal capacity; the relationship of steam to other sources of thermal capacity; and the quantities, costs and related percentages of electricity generated by types and kinds of fuel used, both regionally and provincially."

Jan 1, 1966

By A. G. Scobie

A spectrographic method for the determination of small amounts of gold and the platinum metals in ores has been developed. By means of this method, an assay of 0.001 ounce per ton on one assay ton samples is possible, and by seecrrographic adjustments the sensitivity may be increased to 0.0001 ounce per ton. All the prec10us metals may be determined from one spectrographic arcing with an accuracy of about 5 per cent average deviation from absolute values, and the complete analysis of an ore or similar product may be reported five or six hours after receipt of sample.

Jan 1, 1945

By T. Patching

"Introduction The occurrence of coal was first reported in the Drumheller area in 1793 but it was not until 1912 that production of coal was recorded. In that year 13,000 tonnes were mined in the Drumheller area. Three years later the Drumheller mines ranked next in the Province to those in the Crowsnest and Lethbridge districts, with an output of 225,000 tonnes. By 1917 production had increased to 600,000 tonnes, moving Drumheller to second place in output in the province, second only to the Crowsnest district. It held this rank until 1960. Production, almost all from underground mining, rose to over a million tonnes a year by 1920 and remained between one and two million tonnes per year until 1954. Competition from newly discovered oil and natural gas, beginning around 1950 caused a steady loss of domestic coal markets, especially from underground mines. By 1960 Drumheller's production had fallen below a half a million tonnes per year and the field ranked third in output in the province, being surpassed by new stripping operations in the Castor field. By 1965 only a couple of producing mines remained in the Drumheller valley, and output was only about a quarter of a million tonnes per year(1). Throughout the life of the field production was seasonal, depending mainly on the domestic heating market. In 1979 the last mine ceased production. For the last several years of its operation it had remained open in the summers to provide tours for visitors(2). The Drumheller mining operations were located along or near the Red Deer River valley which is 300 to 400 feet deep, and extended from a few miles above the city of Drumheller to East Coulee, about 13 miles down the river in a south-easterly direction. Intervening mining communities existed for some years at Wayne and Rosedale. During the peak years of production some 2000 to 2500 miners and their families lived in these communities and the surrounding areas."

Jan 1, 1985

By I. H. Warren, F. A. Forward, A. Vizsolyi

"Observations are reported on the pressure leaching of material produced by subjecting chalcopyrite to two types of heat treatment. Calcination of the mineral at 825°C, in the absence of oxygen, induces some sulphur loss by volatization, producing a material from which copper can be leached more rapidly than from chalcopyrite. Treatment of the mineral with elemental sulphur at 475°C leads to sulphur uptake accompanied by a change in mineralogical structure of the chalcopyrite. Copper is prefer-entially and rapidly leached from the sulphur-treated material. A tentative explanation of the sulphur-chalcopyrite re-action is presented."

Jan 1, 1968