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By Joseph N. Bakambu

It is currently recognized by mining society that autonomously guided underground mining vehicles may greatly increase the efficiency of mining production and improve human safety. Recent results in mobile robotics, data fusion and sensor development provide a solid basis for solving the problems related to autonomous underground navigation. Recently developed commercial systems mostly rely on the infrastructure that is expensive and difficult to maintain. Guideless navigation in unstructured and harsh environment of underground mines is still a challenging task. In this paper, we describe an autonomous navigation system for an underground mining vehicle equipped with on-board dead reckoning, inertial and optical sensors. This system does not require any artificial infrastructure and is capable to guide the vehicle along the complex environment of interconnected underground drifts; it also ensures the detection of obstacles on the vehicle path. The user interface for monitoring and control is described and the results of laboratory tests and on-site experiments are presented. Work in progress addresses the system integration for real equipment, and the system upgrade, adding map-building and selflocalization capabilities .

May 1, 2002

By M. D. Quinn

THE sand and gravel used for backfilling at the Hollinger mine is obtained from deposits in a low range of hills some three miles north of the mine, where the Company owns a number of claims. The material is transported to the mine by aerial tram. Shipments from these deposits to the Hollinger mine commenced in June, 1927, and since May, 1936, sand has been supplied also to the McIntyre mine. During 1942, Hollinger requirements were 1,700 tons per day, and 900 tons per day was delivered to McIntyre. For the past seven years production has averaged 900,000 tons annually, and total deliveries to the end of 1942 were 13,129,022 tons, of which 11,075,339 tons went to Hollinger and 2,053,683 tons to McIntyre. Some years ago, when the original pit was approaching exhaustion, new workings were opened on claims east and south of the tramway terminal. It soon became apparent that the sand varies considerably in character from place to place. The weight per cubic foot was found to range from 85 to 130 lb., and the water percolation rate from 2 to 50 inches per hour. As the mine requirement is a sand with percolation rate of 6 inches or better per hour, it was evident that in certain parts of the deposits the material was not suitable for use as backfill. As a consequence, it was decided to carry out a programme of systematic sampling of the claims in order to select the areas containing suitable material for back-fill and at the same time estimate the tonnage available.

Jan 1, 1944

By H. Ranjbar

The Darrehzar porphyry Cu deposit is situated in the Urumieh-Dokhtar magmatic suite of Central Iran. The REE and trace element data suggest that the deposit lies within a continental margin setting. Phyllic and propylitic alteration are pervasive in the area but potassic and argillicalteration are not readily recognized on the surface. This is possibly the result of super gene weathering of the surface rocks or because of an intense phyllic overprint. Poor soil development and abundant outcrop make the area suitable for a case study of lithogeochemical exploration technique for Cu deposits. A total of 262 rock samples was collected over the area and analyzed for 15 elements. The geochemical data cor-relate with the mineralogical composition of the rocks with the phyllic zone being marked by high-er concentrations of Si, Al, K, and by a high sericitization index while Ca, Mg, Sr, total Fe and Mn are more abundant in the propylitic zone. The lithogeochemical study indicates a metal zonation from a Cu + Mo-rich core to a Zn + Pb ± Ag-rich outer zone. Principal component analysis is useful for separating element associations, which in turn reflect the alteration halos. Principal components 1 and 2 depict alteration zoning, where-as 3 shows an association of Cu and Mo, that roughly defines the mineralized area.

Jan 1, 1997

By R. Blosser

Phelps Dodge introduced a process improvement model known as Quest for Zero in 2001. In 2004, the focus of Quest for Zero shifted from corporate-led initiatives to projects led by site personnel. The Hydrometallurgical Division at Morenci focused on two (2) significant process improvements: reducing electrolyte bleed and organic losses. In September 2004, Morenci kicked off a bleed reduction project to reduce electrolyte bleed by 45%. By the end 2005, electrolyte bleed was reduced by 55% through understanding and reduction of variance. Significant process changes made during this time period include changes in acid-addition strategy, reduction of uncontrolled bleed, and optimization of bleed placement. In April 2005, the organic reduction project began with the goal of reducing losses by 15%. Operational improvements during this project included improved organic recovery systems, mixer speed optimization, and picket fence installation. By the end 2005, this project reduced organic losses by 35% and current losses are 48% lower than at the start of the project. These projects and smaller Questfor Zero projects have resulted in sustainable annualized savings of $13 million in Morenci hydrometallurgical operations. Through the use of Six Sigma tools, process variances can be controlled and reduced.

Jan 1, 2007

By G. Grassmück

The question of whether the air curtain principle could be used under conditions prevailing in Canadian metal mines was investigated. A vertical, portable BERRY air curtain was tested, at first in a wind tunnel built at Ecole Polytechnique at full scale (8-by 8-ft cross section) to simulate a drift and its lateral connections, and then in the haulage ways of a metal mine under actual conditions. The results permit us to predict under what conditions this type of air curtain is capable of acting as an air stopping or flow regulator and how it could be further developed.

Jan 1, 1969

By J. P. Grau, J. Z. Lu

"Thickeners/clarifiers are widely used in coal preparation plants, potash processing and other mineral processing facilities. Thickeners play a very important role in slurry treatment systems to accomplish solid/liquid separation. The demand for larger plants and the increased thickener handling capacities sometime require thickener diameters as large as 200m. The development and use of high capacity thickeners makes it possible to increase the handling capacity of these thickeners by several to more than 10 times compared to conventional thickeners. Hence, for a given slurry rate, selection of a high capacity thickener will significantly reduce the footprint of the thickener. Much higher handling rate, higher concentration of underflow, clearer overflow and lower capital cost are obvious advantages of high capacity thickeners; however, some process requirements and the concerns of operational cost, especially the flocculant cost of a high capacity thickener operation, keep some engineers selecting conventional thickeners and staying away from the high capacity thickeners. Forty years after the first high capacity thickener was developed, there are still arguments in the selection of conventional thickeners or high capacity thickeners in mineral processing. The arguments will continue unless there is a clear picture of capital cost vs. operational cost for each of the thickener types. The purpose of this paper is to analyze these costs and to provide guidelines for selection of thickeners. INTRODUCTIONThickeners/clarifiers are widely used in various slurry treatment systems. For mining and mineral processing, thickeners are an integral part of the solid/liquid separation system. The feed to the thickeners are processing slurries. The fundamental mechanism for the solids/liquid separation is for the solid particles in the slurry to gravity settle inside the thickener tank. The clear water overflows from the top of the thickener tank and is typically reused in the processing circuit, and the concentrated underflow is discharged to different downstream process systems depending on the application of the thickeners. The general requirements for a thickener operation include high density of underflow and clear overflow. Since thickener operation consists of continuous feeding, overflow and discharge of underflow, the solid particles will gravity settle against an upward flow with its velocity dependent on the feed rate and the underflow discharge rate. Higher feed rates to a given thickener produce higher upward velocities. Underflow discharge is typically relatively small compared to the amount of overflow. The upward flow tends to take particles up to the overflow. Only those particles having a settling velocity greater than the upward flow velocity will possibly settle out. Hence, thickener feed rate has to be controlled so that the required overflow water clarity can be achieved."

Jan 1, 2012

By C. R. Whittemore

IT can be truly stated that no branch of the engineering industry has progressed more rapidly in the last few years than that of welding. In spite of depressed economic conditions, this industry has continued to advance along scientific lines, replacing the old hit-and-miss method of welding. The development work is intensive, and increased activity is recorded in every sphere of the industry, new applications of existing methods, together with new processes, being added to those already proved in practice. The general acceptance of the welding processes can be measured by their utilization as a manufacturing tool in the metal-working industries. Many manufacturers have partially or wholly adopted welding. These concerns have on their staffs welding specialists, design engineers, and metallurgists, who are qualified to design and supervise the manufacture of welded parts, at a considerable saving to the purchaser. The steady, consistent, and sound growth of welding is based upon investigation of the fundamentals of the various processes, together with their proper application. The classification of welding processes varies in different countries. In Europe, they are divided into two main classes, namely, heterogenous and autogenous. In the former, the parts are joined, as a ruler without melting the edges, and by using as filler a metal different from that of the object welded. Such welding differs from soldering and brazing in that higher temperatures are required to effect the union of the parts. In autogenous welding, the parts are joined by melting of the edges, with or without the addition of a material similar to the parts being joined.

Jan 1, 1936

By C. Pelley, S. Kelebek, T. Katsabanis, B. Nanthakumar

"As part of a larger investigation on generation of fines from mining activities and their effect on downstream operations, ore samples from Onaping area stockpiles of Falconbridge Ltd were studied. Batch flotation tests were conducted on the stockpile samples and fresh ore samples to explore the effect of stockpiling on grade-recovery performance. Head grade and particle size distribution of the samples from stockpiles were adjusted so that the samples compared are similar to each other. In additional tests, fresh ore samples were blended with stockpile samples at various percentages, representing 27.4, 50.2 and 81.5% of stockpile samples.The samples from stockpiles produced inferior nickel grade-recovery curves compared to fresh ore. The poorer grade-recovery performance was attributable to two factors. First, a significant dilution effect was observed due to the nearly complete flotation of pyrrhotite. Secondly, while pyrrhotite from stockpiled ore became highly floatable, the same stockpiling conditions caused nickel losses that were attributable to non-floatability of some pentlandite.In the case of blending, there was generally insignificant difference in the nickel grade-recovery between fresh and blended samples when the blending cases involved 27.4 and 50.2% stockpile samples. However, an increase in pyrrhotite recovery was observed for both blending cases. At 81.5% blending case, degraded grade-recovery was significant primarily due to the nearly complete flotation of pyrrhotite. There was a small, but noticeable amount of nickel loss for this case. However, for lower ratios of blending this loss was negligible.The increase in pyrrhotite recovery was partly attributed to its activation by Cu and Ni ions. Formation of a significant amount of elemental sulphur was also apparent from analysis of stockpiled samples. It can be concluded that pyrrhotite is mildly oxidized under stockpiling conditions while part of the pentlandite undergoes a more severe oxidation."

Jan 1, 2006

By V. N. Bryant

"1.0 SYNOPSISThe Colomac property, located 137 miles by air northwest of Yellowknife in Canada's NWT. will recover gold from 10,000 tons per day of open pit ore.The decision to proceed with the project was based on a mineable ore reserve of 16 million tons (including dilution) averaging 0.064 oz/ton. After an in-fill drilling program in the latter half of 1988, Colomac's mineable, proven and probable ore reserves were increased to 28 million tons averaging 0.056 oz/ton of gold.A feasibility study completed by Wright Engineers for Neptune Resources in December 1987 examined the possibility of vat leaching followed by C.l.P. and electrowinning. A second feasibility study completed by Wright in February 1988 demonstrated that a conventional agitated leach circuit using a finer grind justified the higher capital cost of this process. This study was accepted and formed the basis of the project authorization.Preliminary engineering commenced in April 1988 and detailed engineering was started in September of that year with the objective of a mine/mill start-up by early 1990.The short time period available to meet the project deadline was additionally challenging, as site access is either by plane year round, and winter road from early February to late March. Normal lead times for equipment design, specifications, quotations and bid analyses were not possible, if the winter road time-slot was to be met."

Jan 1, 1990

By M. Lindgren, M. Latva-Kokko

Process performance can be significantly influenced by proper design of hydrometallurgical reactors. Optimal reactor configuration is case specific and depends on process requirements such as reaction rate limiting factor and particle size distribution. Choice of materials and manufacturing methods also play important roles in highly corrosive and abrasive environment often involved in hydrometallurgical industry. In this paper performance of different reactor designs has been compared in some typical hydrometallurgical applications. Results show that significant improvement in energy consumption, gas utilization efficiency and equipment lifetime can be achieved with process specific reactor design.

Jan 1, 2014

By George C. Coupland

"THE PRELIMINARY EVALUATION of the orebodies was computerized for tonnages and grades on a 20- by 20-ft grid. The evaluation was based on drill-core intersections over 40-ft vertical projections coinciding with the proposed bench intervals, taking the deviations of the drill holes into account. The computer print-outs gave an ore distribution and evaluation bench by bench and indicated the areas of varying mineralization. The two orebodies have a combined length of 2,200 ft. and a maximum width of 600 ft. They strike almost due north, dip about 85 degrees east and rake to the north. The computer print-outs were used to design an open pit with an over-all waste to ore ratio of 2.58:1. At the outset, the waste to ore ratio was 4:1. The pit is tear-drop shaped, with the narrow end to the south, and at the surface is 2,600 ft long by 1,800 ft wide. Ramps are 100 ft wide, with a 10% grade. The south ramp is left-handed, starts at the southeast and will continue to the bottom of the pit. A second ramp on the west and north walls runs from surface to join the south ramp at 7 bench. This ramp will handle the total pit production when part of the south ramp, from 4 bench to surface, is lost due to the underground operation. Over-all wall slopes are 53 degrees, and are laid out to avoid prominent protuberances which could relax and become unstable and potentially dangerous (Fig. 1) ."

Jan 1, 1974

By M. Hitch, M. B. Lytle

Resource development conflicts arise largely due to differences in worldview between those who support and those who oppose development. This paper illustrates how societies and civilizations with different worldviews produce different levels of creative and economic energy. A questionnaire was created to build a database of responses to test whether worldview and resource development are correlated. Factor and multivariate regression analysis determined correlations for respondent groupings. Results showed a positive correlation between the respondents\u2019 worldviews and their attitudes toward resource development. Correlations became stronger as the homogeneity of the sample subgroups increased.

Jan 1, 2017

By X. M. Zhou, T. Wu, He X. N.

"In order to study the support effect of concrete in the surrounding rock of a metal mine, the concrete and surrounding rock were comparatively studied. Firstly, the general situation of the ultra-kilometer deep shaft in Xincheng, surrounding rock and geological conditions are introduced. Secondly, the experimental scheme of material mechanical properties of surrounding rock and shaft lining concrete based on same standard dimension and shape are examined, and the mechanical tests of rock and concrete based on different shapes and sizes are carried out. The study shows that the compressive strength value and elastic modulus of the rock cube are 1.28 times and 1.78 times that of the cylinder, and the Poisson's ratio is 0.77 times. The dimensional effect law of the concrete cube and cylinder test data is consistent with the rock law. However, the test results of the concrete are less discrete than the rock; the test also shows that the strength and elastic modulus of the concrete lining of the main shaft of the Xincheng are much smaller than the surrounding rock that is also used as the concrete aggregate, which indicates that the support of concrete lining in the surrounding rock shaft behaved weakly. Based on the development of the modern NATM rock support theory, it is suggested that the mechanical test standard of artificial support materials should be unified into the mechanics test standard of natural rock, which will help to improve the accuracy and practicality of scientific calculation of geotechnical engineering.INTRODUCTIONConcrete and rock mass are usually attributed to a class of materials called frictional materials or geotechnical materials, of these two kinds of materials, the first is man-maded and the second is natural. The attentive scholar can find that there are many differences in the test criteria for the mechanical properties of the two kinds of materials among different industries or departments, reflecting the differences in their sources, tradition, and the application thoery of material strength among them.There is a situation wordwide, that there are many testing standards of the same material mechanical parameters such as elastic modulus, compressive strength, etc. The cube with a side length of 150 mm is adopted as a common sample in the architecture department in China, while a cylinder with a diameter of 50 mm and a height of 100 mm is as a standard specimen in geotechnical department.Knowing that mechanical parameter values are strongly affected by the the size, shape of specimen and loading rate etc, that is to say the mechanical parameter values of the same material are completely different when obtained from different test with different standards. However, such a problem has been long neglected in geotechnical engineering."

Jan 1, 2019

By Sanjiv Dhanjal

"Often mineralogy dictates the physical properties of a material rather than chemistry. This in turn affects how that material behaves in a process plant, and whether optimum conditions for refining one mineral are different to those for refining another. For example how does refining copper from chalcopyrite differ to refining copper from bornite?Minerals are naturally occurring inorganic solids that have a definite chemical composition and an ordered atomic (crystalline) structure. X-ray diffraction is a common and highly developed method to identify and quantify minerals. X-rays are fired onto a material and the mineral is identified by the angles where the diffracted X-rays are in phase.Most X-ray analyzers are laboratory based. A sample is prepared and analyzed by a detector that travels in an arc to collect the diffraction pattern. COSMA has a curved detector which picks up the complete x-ray pattern instantaneously. Therefore, material is collected and transported through the machine where is it is analyzed continuously.COSMA measures about 0.5 tons of material per day, while laboratory analyzers measure a few grams at a time. Real time data trends are produced allowing for improved process evaluation and automatic control. The COSMA on-line system has taken X-ray diffraction out of the laboratory into the field.There are a number of systems in operation in cement production. It is recognized that on-line mineral analysis has a role in many other processes. This paper presents the technology and invites further discussion should the technology be beneficial to members of the audience.INTRODUCTIONThis article describes COSMA, a system that provides direct and continuous on-line mineral analysis using X-ray diffraction in the field. While X-ray diffraction is a common technique used in mineral analysis, the time from sampling to analysis can be hours and plant operators typically do not get the data in time to control a process or respond to an upset. COSMA provides rapid measurement, analysis and feedback of data, in real time. This provides powerful information that can be used to optimize operation, thereby reducing costs and increasing quality."

Jan 1, 2009

By T. H. Janes

"IntroductionSULPHUR is a constituent of many industrial products. It is part •of almost anything we eat, wear, or use. In fact, sulphur is so important that there has been an increasing tendency in recent years to measure industrial growth in terms of sulphur consumption.The uncertainty of obtaining adequate supplies •of elemental sulphur from the Frasch-mined domes. of Texas and Louisiana in the United States led all countries in 1950 to seek alternative sources of supply of this necessary element.Canadian industry has been supplied, for many years, with most of its sulphur requirements by United States Frasch-mined sulphur. World demand for this sulphur has necessitated the imposition of controls on its distribution, as production has not kept pace with demand. Until the third quarter of 1950, Canadian manufacturers of pulp and paper, heavy chemicals, explosives, and textiles •experienced little difficulty in obtaining their needs of elemental sulphur from the United States. However, at that time, low producers' stocks and declining reserves threatened to reduce shipments from producers, with the prospect of a continuing shortage for some time. These conditions, together with higher prices, stimulated the search for alternative sources of supply in Canada."

Jan 1, 1952

By G. Warren

"In the past, smelting and refining of the precious metals recovered from ore has been generally looked upon as a small and relatively minor component of the business of gold mining. This is not surprising when one considers that the manpower, space and capital components of a gold refinery at an operating mine can be several orders of magnitude smaller than other unit operations at the same location. This philosophy is changing however, as economic and environmental pressure are forcing a re-evaluation of all aspects of gold mining and milling.This paper describes and evaluates three furnace technologies available and used by precious metal operators. The operation of arc, induction, and reverberatory furnaces are discussed and their strengths and weaknesses with respect to capital cost, impurity removal, slag quality, and environmental concerns highlighted.INTRODUCTIONIn the past, smelting and refining of precious metals recovered from ore has been generally looked upon as a small and relatively minor component of the business of gold mining. This is not surprising when one considers that the manpower, space, and capital components of a gold refinery at an operating mine can be several orders of magnitude smaller than other unit operations at the same location. Generally, gold producers will spend a great deal of effort to squeeze the last drop of efficiency out of a leach or grinding circuit via simulation or testwork but will design and build the refinery based upon past practices or industry standards. That past practices or industry standards will provide a functional refinery is undisputed; that this philosophy achieves the efficiencies that are possible or demanded in other plant operations is debatable.Environmental and worker hygiene requirements are considerations as well. Pollution control equipment can become a substantial contributor to both the capital and operating costs of a refinery and furnace selection has a marked influence upon the type and cost of this equipment. Slag characteristics and subsequent processing options available also play a role in furnace selection as precious metal slag losses and the recovery effort can greatly influence the cost / benefit analysis of a particular technology."

Jan 1, 2001

By J. A. Brusset

Introduction I am indebted for most of the information in this paper to Mr. P. S. Lea, Deputy Chief Inspector of Mines of Great Britain, and Mr. A. E. Crook, Engineering Inspector of Mines, Ministry of Fuel and Power. Other coal-mining men in Canada may be as surprised as I was to learn that there are twenty-five Diesel locomotives in service underground in British gassy coal mines. Many more are in use underground in non-gassy coal mines and in metal mines. Approval has been granted for an additional fifty flame-proof locomotives to be used in gassy mines, and these should be in operation by the end of 1946, or as soon as manufacturing facilities permit.

Jan 1, 1946

"In mining operations, it is widely accepted that strategies to drive operational goals must revolve around improving plant availability, reliability and process efficiency. In this paper, an integrated solution framework for reducing process variability, increasing process efficiencies and maximizing plant availability and reliability is discussed within the context of a typical milling circuit. The framework exploits a real-time analysis “gap” evident in production operations that prevents timeous identification of equipment and process failures. This is achieved by leveraging predictive analytics for equipment health monitoring, control loop monitoring and advanced regulatory control solutions towards enhanced process performance, improved asset management and regulatory compliance. The framework provides for early and actionable warnings on equipment and process degradation using real-time operational data.INTRODUCTION The changing landscape of gold mining and processing presents many challenges that threaten sustainability of existing (marginal) operations as well as deter investment in new projects. As existing “shallow” ore resources become depleted, mining operations are increasingly treating low-grade ores with complex and randomly variable mineralogical characteristics. For example, in South Africa, historically one of the major gold producers, the average grade of ore treated has been decreasing from averages above 4.5 – 5 g/t to current grades averaging less than 3 g/t within a decade (Chamber of Mines, 2012). Moreover, lower tonnages are being treated as a result of the narrowing of the available resource base. Also, the regulatory framework has tightened over the years due to heightened concerns on the safety, health and environmental impacts of mining and related activities."

Jan 1, 2014

By Craig M. Sams

"Variable speed mill popularity has grown during the last 10 years with almost every SAG mill fitted with a variable speed drive. In addition, many of today’s larger ball mills are running a variable speed drive for various reasons. In the past, motor suppliers, especially gearless drive suppliers, have considered mills as being an overpriced rotor. This has lead to mills demanding more torque than what a motor can generate. This paper identifies the reasoning of what has happened and how to mitigate the mismatch in the future. Technical and economical issues are discussed, allowing mill operators to make an informed decision of which path to take in deciding what motor to put on a mill.INTRODUCTIONIt has been generally accepted in the past that motor suppliers and specialists in engineering companies understand the workings of a SAG mill or a ball mill and although there is a fair knowledge within the industry, in regards to mill torque requirements, the advent of variable speed drives in ball mills has shown the writer that a complete understanding of the mill torque requirements are not fully understood.This paper will discuss two aspects of mill torque requirements, (operating torque at various charge levels and start up torque), and how it relates to specifying the proper motor torque requirements.MILL OPERATING TORQUEMill operating torque at a given charge level and density in a mill is essentially constant throughout the entire speed range. For the purpose of this paper, we will take a 38’x 22.5’, 20’ EGL, SAG mill and provide the power draw at variable speeds but fixed loading conditions.The power draw and mill torque of this mill at 8% ball charge and 28% total loading (Highlighted) is shown below as Figure 1, at variable speeds."

Jan 1, 2007

By O. Peters, D. Lascelles, R. Caldwell

Ageing tests on a crushed Cu-porphyry ore were conducted to assess the impact of sample oxidation on flotation performance. The objective of the testwork was to simulate ageing of ore that would be blasted through block caving but not milled for possibly several months. Ageing was performed in specially designed columns, using site and deionised water to keep the samples moist over the entire ageing period. Samples from each ageing interval (0, 0.5, 1, 2, 4, and 8 months) were examined for physical and metallurgical changes resulting from the ageing process through chemical analyses, size analyses, and batch flotation tests.

Jan 1, 2009