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By J Zhou

Based on the mineralogical characteristics and mineral processing techniques required, gold ores can be classified into 12 types, including gold placers, quartz vein-lode ores, oxidized ores, silver-rich ores, iron oxide copper gold ores, porphyry copper gold ores, iron sulphide ores, arsenic sulphide ores, antimony sulphide ores, bismuth sulphide ores, telluride ores and carbonaceous sulphide ores. The bulk mineralogy and gold deportment of each ore type is different and may have a significant impact on ore processing and metal extraction. This paper summarizes the bulk mineralogy and gold deportment in some major types of gold ores and discusses the potential impact of gold deportment on gold metallurgy.

Jan 1, 2011

By Côme Carbonneau

The ten -year case history of La Société Québécoise d'Exploration Minière (SOQUEM) is reviewed under the headings of the company's raison d'être, objectives, policies, structure and final goal. A brief analysis of its past operations is presented: aborted efforts, findings and developments. A projection is made of the future earnings, in terms of estimated present value, of its three main discoveries - Louvem , Niobec and Selim.

Jan 1, 1978

By R. H. Batey

"The process control and cost of monitoring mining slurries is normally accomplished using a continuous mass flow system, conventionally embodying a volumetric magnetic flow meter and nuclear density meter. DuPont’s Maxville Mine at Starke, Florida, USA was chosen to examine costs associated with performance of a nuclear density meter. Since it is typical of many mid-sized establishments and has a long history in their use. The published accuracy under ideal conditions of a nuclear density meter is ± 1.0%, but in practice considerable ambiguity of cost in annual production implies ± 2% at best. Using such accuracy in a 400mm (16”) pipe carrying zircon slurry at 75,000 tons/day, the cost of ambiguity, together with initial and running costs, amount to nominally $962,000 per year.Such cost leaves much to be desired. This paper describes an unique continuous, direct mass per unit volume Density Meter used for mining slurries in pipe sizes up to 1000mm (40”), which significantly reduces the cost of ambiguity of measurement and running costs. At the same time the disadvantages of nuclear devices are obviated.In an effort to reduce costs, a 400mm non-nuclear Density Meter was installed at the DuPont Maxville mine. A typical on-site accuracy can be shown to be ± 0.285%, whereby the cost of ambiguity, initial and running costs in the example above shows an annual improvement of $78,660, a reduction of over $880,000 per year.Even higher cost savings may be demonstrated with mass flow systems having gas entrainment compensation. Unlike nuclear and other inferential density meters, the signal quality of this non-nuclear Density Meter is totally unaffected by entrained gas found in mining slurries. This non-nuclear density meter is then ideally used with a vortex velocity domain sonic flow meter, and provides for the first time compensation in wet density or dry % solids, as well as wet or dry mass flow of mining slurries with entrained gas, resulting in enormous reductions in cost ambiguity."

Jan 1, 2013

By Gerardo Vargas

Without an adequate data sample there can be no Reliability Analysis (RA). Without analysis there can be no systematic, verifiable improvement in reliability or in operational economy. A sample is a collection of life cycles. A beginning and ending event define a failure mode life cycle. Well known obstacles impede the reliability engineer in his role to analyze maintenance data. The main problem lies in the difficulty to obtain analyzable data samples. Although modern Computerized Maintenance Management Systems (CMMSs) should provide the needed information they rarely deliver samples of adequate quality. Cerrejon, an integrated mining operation in Colombia, South America solved the information management problem by applying a "Living Reliability Centered Maintenance (LRCM)" process to its fleets of Trucks, Graders, Dozers, Loaders, and Shovels. This paper describes a method wherein completed work orders capture reliability analysis enabling information. The 'right' maintenance observations reference significant failure modes. Work orders link to records in a continuously growing Reliability Centered Maintenance (RCM) structured knowledge base. Grouped and filtered knowledge-to-work order links, being instances of failure modes, provide the samples required by reliability analysis software tools and methods. LRCM software and related procedures facilitate the growth of knowledge, manage the work order-to-RCM relationship, and generate samples for subsequent reliability analysis.

May 1, 2011

By G. D. West, J. D. Purdue

"Marked Ball Wear Testing (MBWT) employs identified grinding balls which are placed in an operating mill for a period of time, retrieved from the mill, and then measured to determine the amount of wear experienced. This is an excellent method to evaluate the relative differences in the wear rates of various grinding ball materials and/ or manufacturing process variables simultaneously and under actual production milling conditions. This paper outlines the methods of ball identification, testing and evaluation procedures and discusses the strengths, weaknesses, and proper application of marked Ball Wear Test results. INTRODUCTIONThe consumption of grinding media in the mineral processing industry represents a substantial cost to the comminution process. Grinding media is most often found to be the second highest consumable cost next to energy. To determine the selection of the most cost effective media, one must determine the specific wear mechanisms present in a given grinding application and decide which media system addresses these mechanisms most economically.A MBWT has definite advantages over other methods used to evaluate the relative consumption rates of grinding media. It allows the comparative testing of a large number of different ball types under identical operating conditions in a very short period of time. In addition, production wear rates can be accurately predicted directly from MBWT results. Typical MBWT test periods are in the order of 350 to 700 operating hours for high wear rate application and upwards of 2000 or 3000 hours for low wear rate applications. These times are much less than the 5000 or more hours required to purge a mill of the incumbent ball charge (to convert it to the test ball) and then provide even a minimal test period in which to conduct a full production test, for just one type of ball. Under production test conditions the evaluation of competitive product or the development of improved grinding balls is necessarily slow."

Jan 1, 1991

By M. Nusheh, C. Bumby, A-H. Puichaud

"Vanadium-bearing titanomagnetite ironsand deposits occur along a 480 km stretch of the west coast of New Zealand. In the present study, the leaching behaviour of NZ ironsand was investigated in hydrochloric acid (2–10 M) solution at 50–80°C for 2 h. The resulting leaching liquors exhibited Fe/V ratios in the range 150–180. The distribution of trace element contaminants between solution and solid residue was investigated, and processing conditions have been determined for rapid and efficient dissolution of Fe, Ti and V into the leaching solution. In addition, mineralogical analysis of raw ironsand and leached residues was performed in order to study the leaching mechanism. Samples were characterized by scanning electron microscopy, energy dispersive X-ray spectroscopy (EDX), X-ray fluorescence spectroscopy, qualitative and quantitative X-ray diffraction and inductively coupled plasma mass spectrometry (ICP). Results indicate that the dominant dissolution process follows shrinking particle mechanism. However, ironsand particles containing ex-solution lamellae exhibited a different leaching behaviour compared to homogenous titanomagnetite particles.INTRODUCTION Titanomagnetite ironsand (blacksand) forms Quaternary onshore beach and dune deposits and offshore marine deposits (Figure 1) along 480 km of coastline from Kaipara Harbour south to Wanganui on the west coast of the North Island (Christie, 2016). It attracted the attention of scientists from the earliest times (Galbreath, 1998). The principal minerals in ironsand are titanomagnetite (TTM), ilmenite, quartz and ferromagnesian minerals such as augite, hornblende, diopside, and olivine with smaller amounts of rutile and apatite (Mauk, 2016).Several large ironsand deposits have been discovered and exploited since the late 1960’s. The Waikato North Head and Taharoa mines are currently operating and they have a 2015 measured resource of 400 million tonnes in total and provide over 2 million tonnes of concentrate per year by simple cost effective magnetic separation process (Christie, 2016; Christie & Brathwaite, 1997). The concentrate contains not only 7–8% TiO2 and 56–60% SFe, but also 0.3–0.6% V2O5, is both a valuable titanium mineral and a source of vanadium. All the production is used for steelmaking industries. The other potential for utilizing more of this resource without requirement of integration into an iron and steel producing plant is to develop a hydrometallurgical route with advantages of higher efficiency along with lower emissions. Currently, there is no existing commercial process capable of extracting all three of these elements as industrial-grade commodity products. Acid leaching of TTM concentrate with higher degrees of recovery and selectivity is a key step in any possible hydrometallurgical route."

Jan 1, 2017

By Kenji Koga

In order to derive the population distribution of gold nanoparticle morphologies as a function of size, highresolution electron microscopic (HREM) observations and structural analyses were done for over 4000 particles (3-18 nm) grown by an inert gas aggregation (IGA) source using helium. To obtain structural features threedimensionally and not to skip high-contrast images, twenty photos were taken for each particle by changing tilting angle of a specimen holder in combination with the through-focus procedure. For the reliable structural determinations, HREM images and their tilting sequences were directly compared with multislice-simulation results for icosahedral, decahedral and face-centered cubic morphologies, and finally the population distribution of each morphology was obtained as a function of size. In the present sample, icosahedral particles were dominantly populated over the whole observed size range and decahedral particles with non-spherical truncated bi-pyramid shape were secondly dominant. Even though a large number of sampling has been done, few crystalline particles were observed.

Oct 1, 2003

By Peter Price

Introduction The Horne mine, the property of Noranda Mines, Limited, has in the last six years become one of Canada's leading gold and copper producers. Since the smelter was blown-in during December, 1927, up to the end of 1933, it has produced approximately 1,200,000 oz. of gold and 35O,OOO,OOO lb. of copper. The property is situated within an area occupied entirely by pre-Cambrian rocks and characterized by the complexity of rock types and structures common in such terrain. But apart from this the mine also presents other features of unusual interest in the size, number, and variety of the ore-bodies, the mineralogical associations, and the rather exceptional values both in gold and copper. It is not surprising, therefore, that a study of the geology of the mine should bring to light many interesting problems. The present paper summarizes, as concisely as possible, the results of such a study. Actually, it is a much abridged version of a more comprehensive and detailed paper completed by the writer a year ago (1), but embodying certain corrections and additional data made possible as a result of further development at the mine and of field work carried out during the summer of 1933. Owing to limitations of space, the paper is, in the main, a presentation of conclusions arrived at in the course of the study. Much of the supporting evidence cited in the thesis referred to has, of necessity, been omitted here.

Jan 1, 1934

By Robert B. Parsons

"Canada's mining industry has reason to celebrate the federal government's budget which was tabled on February 26, 1986. The budget contains an important proposal to improve the flowthrough share rules by effectively eliminating the investor's liability to third parties. This proposed rule change and other important budget proposals are the subject of this month's edition of Tax Notes. IntroductionWhile Finance Minister Michael Wilson's February 26, 1986 budget received a mixed reaction from business and labour leaders, the budget was mainly ""good news"" for the mining industry. A long-standing bugaboo in the flow-through share rules will be cleared up, a permanent set of rules will be installed for the tax treatment of Northern allowances, and limited partnerships will apparently be able to continue to serve a useful role in financing exploration programs. Flow-Through SharesThe present rules which apply to an investment in flow-through shares are structured in such a way that the relevant Canadian exploration expense (""CEE"") must be incurred by the investor, and not by the issuing corporation. Obviously, in the case of most flow-through share issues, this requirement that the investor incur the CEE poses a practical problem-the investor typically does not have the expertise or organization to incur the CEE. As a result, an integral part of most flowthrough arrangements is an agreement whereby the investor appoints another party (usually the issuing corporation) as his agent to spend the flow-through funds on CEE on his behalf. Although the issuing corporation might consequently be the party which actually spends the flow-through share funds on CEE, nevertheless the legal structure is such that the investor is the ""principal"" who is incurring the CEE. As such, the investor is legally responsible for any third-party claims with respect to the exploration program. Third-party liability might arise, for example, in the case of property damage, personal injury, environmental infractions, and so on."

Jan 1, 1986

By B. K. Ayres

Modern management of uranium mine or mill waste storage facilities (WSFs) often requires that they be remediated or decommissioned with a cover system. In addition to creating a self-sustaining landscape that supports the end land use, cover systems are intended to reduce long-term risks to human and ecological receptors from the stored waste to acceptable levels. The performance and longevity of a cover system will be strongly influenced by the features or shapes of the WSF final landform. Hence, cover systems and landforms must be thought of as an integrated design. Based on the author’s experience, the most important elements of the design process for maximizing the potential of WSF final landforms / cover systems being sustainable over the long term are noted in this manuscript. Several key items that must be considered and addressed during the planning phase of constructing a WSF final landform are also described. Various methods for monitoring performance of a WSF final landform are also reviewed. Key aspects associated with decommissioning a tailings WSF and waste rock pile at the former Cluff Lake uranium mine in northern Saskatchewan are included to illustrate many of the practical insights provided in this paper.

Jan 1, 2020

By H. Harzer

Statistics Canada record s indicate that in recent years 70 to 80% of the entire Canadian coal production originated from Western Canadian mines, in which about 55 to 65% of the coals are bituminous. Until recently, less than 20% of the latter were extracted by underground methods , the remainder being surface-mined. However, it is estimated that within the foreseeable future up to 80 or 90% will have to be gained by underground techniques. Although Western Canadian coal resources are considerable, they are unfortunately locked in severely distorted, steeply pitching and faulted seams of variable thickness. Consequently, their economic extraction poses serious engineering problems. One possible solution to the problems is to apply hydromechanical methods to underground mining. This is a well-proven method shown to be economically viable under Western Canadian conditions. This paper endeavours to widen the scope of published data on hydraulic coal mining and is based on practical engineering and cost data gathered through contacts with West German designers and operators of hydraulic mines.

Jan 1, 1978

By Owen Matthews

THE property of Cochenour Willans Gold Mines, Limited, is in Dome township, Red Lake area, northwestern Ontario, a hundred miles north of Kenora and close to the Manitoba border. 'Planes of Canadian Pacific Air Lines, Limited, serve the district daily, except Sundays, from Kenora, Ontario, and Lac du Bonnet, Manitoba, passengers being trans-ported from Winnipeg to Lac du Bonnet by car. Freight is handled by the Patricia Transportation Company. It is brought in from Hudson, Ontario, on the Canadian National railway, in summer by water and in winter by tractor trains over the frozen lakes. The first staking on the property was done during the 'rush' of 1926 and 1927 by William M. Cochenour, and on April 30th, 1928, a syndicate was formed, taking in claims of W. M. Cochenour and Dan Willans and H. G. Young. Surface trenching was carried on almost continuously until 1934. In October, 1932, Ventures Limited took an option on the property, and in December of that year they carried out an extensive diamond drilling programme, but considering the results did not justify further work, they relinquished the option. In July, 1934, further surface work was carried out under the supervision of W. P. Mackle and, on the strength of his report, Noah Timmins Incorporated took an option on the property and agreed to make an initial expenditure of $100,000 on development work. This was carried out under the direction of Hollinger Consolidated Gold Mines and included the sinking of a shaft to a depth of 165 feet, establishment of a level at the 150-foot horizon, and some 1,900 lineal feet of drifting and cross-cutting. The results obtained from this work were not considered of sufficient promise and the option was terminated in August, 1935. Mr. Cochenour then approached Alex Mackenzie and associates, and the present Company, Cochenour Willans Gold Mines, Limited, was organized on April 9th, 1936, and sufficient money was raised to continue operations. Under the direction of W. P. Mackle, the shaft was deepened and a second level was established at the 275-foot horizon; but, due to several adverse reports by outside engineers, it was found impossible to raise sufficient money to carry on the work and operations were suspended in December, 1937. In January, 1939, an agreement was entered into with Gold Eagle Gold Mines for the milling in their plant of some Cochenour ore. A total of 4,959 tons was treated, and the results were so encouraging that further finances were raised by C. A. Gentles & Company, of Toronto. Following further development in the mine, it was decided to erect a mill. Construction was started on September 1st and the mill was turned over on November 29th. The first gold brick was poured on December 23rd.

Jan 1, 1942

By Richard W. Hutchinson, Robert W. Hodder

"Pyritic massive sulphide deposits lie conformably above submarine volcanic rocks, and beneath siliceous and manganiferous iron formations. The massive bodies are often zoned, with a barren pyritic top grading downward into pyrite with sphalerite, and occasionally galena, and a chalcopyrite-rich base. They are frequently brecciated, like the lavas beneath them, which contain pyrite and chalcopyrite as coatings on joints and fractures and as dispersed grains. The lavas also exhibit propylitic, argillic and silicic alteration. These sulphide deposits appear to have formed on sea bottom from fluids escaping through the lavas during periods of exhalative and sometimes explosive volcanism. The process may be comparable to the present deposition of base metals from hot brines in the Red Sea. Some porphyry copper deposits resemble the altered, sulphide-impregnated lavas which lie beneath bodies of pyritic massive sulphide. In porphyry copper deposits, pyrite and chalcopyrite occur as veinlets and dispersed grains in intensely fractured tuffs, flows, sub-volcanic intrusions or other host rocks, characterized by propylitic, argillic and silicic alteration. Metal zoning is present; sphalerite and galena often occur in manganiferous veins peripheral to a shell of pyritic veinlets which encloses the mineable zone of chalcopyrite-bearing veinlets. This sulphide distribution appears to have formed from hydrothermal fluids escaping, either passively or violently, from solidifying magma. The fluids may be comparable to hot brines encountered by drilling in the Salton Sea area of California.There are also important differences between massive sulphide and porphyry copper deposits. The former usually occur in differentiated, mafic-to-felsic subaqueous volcanic rocks, whereas porphyry copper deposits are often associated with shallow plutons and subaerial volcanic rocks of intermediate to felsic composition. Calcic-sodic metasomatism accompanies the massive sulphides, whereas potassic metasomatism characterizes the porphyry cop.per deposits. Further, ages of massive sulphides range from Archean (Keewatin) to Tertiary, but porphyry coppers are Triassic to Tertiary in age. These similarities and differences suggest that massive sulphide and porphyry copper deposits are related tectonically and metallogenetically. The former were generated during early stages of crustal evolution, particularly in Archean time, and were repeatedly derived during early stages of younger, individual tectonic cycles. In contrast, the porphyry coppers developed during the later stages of crustal evolution - later in terms of both absolute age and relative age within younger individual tectonic cycles. They appear in younger rocks generally, and specifically in rocks involved in events of the middle and late tectonic stages of orogenic belts. The massive sulphide deposits are of early-magmatic eugeosynclinal affiliation, whereas the porphyry deposits formed in late-magmatic, predominantly subvolcanic settings at the edge of continental plates, where contemporaneous exhalative volcanism was mainly a subaerial phenomenon. Transitional types of deposits exist and are found on continental margins with flanking eugeosynclinal sequences."

Jan 1, 1972

By S. Kelebek, C. Gibson, M. Aghamirian

"Global niobium production is currently dominated by three operating mines, Araxá and Catalão in Brazil, and Niobec in Canada. Although over 90% of the world’s niobium is mined and processed in Brazil, a number of high grade niobium deposits exist worldwide. The advancement of these deposits depends largely on the development of operable beneficiation flowsheets. Pyrochlore, the primary niobium bearing mineral, is typically upgraded by flotation with amine collectors at acidic pH, and often includes a reverse gangue flotation stage prior to pyrochlore flotation. Challenges related to the flotation of niobium ores include complicated flowsheets which require pulp pH adjustment from alkaline to highly acidic conditions, difficulty recovering fine grained pyrochlore, producing concentrates that meet specifications, and minimizing the losses to reverse pre-flotation gangue concentrates, slimes, and tailings. The specific flotation challenges associated with different niobium ores are discussed. In some cases, the use of alternate flotation reagents and flowsheets are required to achieve reasonable gangue rejection and maximize recovery, particularly of very fine pyrochlore grains.INTRODUCTIONNiobium is a transition metal primarily found in the oxide minerals pyrochlore (Ca,Na)2Nb2O6(OH,F) and columbite (Fe,Mn)Nb2O6. After physical upgrading, pyrochlore is converted into a ferroniobium product which is largely used in the production of high strength low-alloy steel. Today’s niobium market is dominated by the Companhia Brasileira de Metalurgia e Mineração, whose Araxá operation produces over 80% of the world’s niobium. There are two smaller operations, Niobec in Canada, and Catalão in Brazil, which account for the remaining production (U.S. Geological Survey, 2013).A number of niobium deposits exist globally, their advancement dependent on the development of operable beneficiation flowsheets. Pyrochlore, the predominant niobium bearing mineral, is commonly upgraded by flotation with cationic amine collectors in acidic pulp conditions. However, many niobium ores do not respond to this typical flowsheet because they are either highly refractory or niobium deports to other minerals such as columbite, ilmenite and rutile, which are not as responsive to this reagent scheme (Bulatovic, 2010). Other reagents, such as alkyl hydroxamic acid, have been studied extensively and while promising, are not without their own inherent issues including lack of selectivity over iron oxide minerals, and high addition rates (Ni & Liu, 2012; Gorlovskii, Eropkiny, Koval, Strelstin, Khobotova, & Shtchukina, 1968; Gorlovskii, Khobatova, & Shtchukina, 1966).It is the aim of this report to highlight the challenges associated with the physical upgrading of niobium minerals through examination of the flotation behavior of two different ore systems. In order to overcome these challenges it is important to thoroughly understand the limitations of typical flowsheets/reagent schemes currently employed in operation."

Jan 1, 2015

By R. C. A. Thurston

"The lack of notch ductility in structural steel in general, and in ship plate in particular, has been the object of widespread attention in the metallurgical world since the spectacular collapse of the ill-fated S.S. Schenectady in 1943. Numerous investigations have been carried out, many new tests have been developed, and a vast amount of useful data has been published, with the result that our knowledge is now far more advanced than it was at the time of this disaster. Interest has focused largely on the medium strength steels, and the various factors which cause a normally ductile steel to behave in a brittle manner in the presence of a notch and at relatively .low temperatures have been thoroughly studied.The literature available on high-tensile steels, however, is less extensive and the present paper is intended to supplement this information. Notch toughness tests have been carried out, at the request of the R.C.N., on a high-tensile manganese steel in the form of commercial ship plate. One of the newer tests, the tear test, and the Charpy V-notch impact test, were used to examine the characteristics of plate for the construction of an ice-breaker. About 250 plates, mainly in the thickness range from 11,4 in. to 1 % in., were tested. An an-alysis of the results obtained is presented, certain aspects of the tear test are discussed, and the correlation between the two types of test is studied. On the basis of information available from casualty ships, the plates, with few exceptions, were considered to be satisfactory."

Jan 1, 1959

By William M. (Bill) Shaver

"The author will review the early history of shaft sinking as well as focus on the improvements made in shaft design and construction over the last fifty years (1970 to 2019) to predict how shaft sinking will change over the next fifty years and describe what the shaft sinking world might look like in 2070. INTRODUCTION It is a great honour to be one of your keynote speakers for the Fourth International Conference on Shaft Design and Construction being held in Toronto in November 2019. This is one of the opportunities we have to thank the Mining Companies of the world as well as the industries and companies we serve for their support of the mining contracting and shaft sinking business of Canada and the world at large. Many of these companies are represented here today. Without your support many of the innovations we will talk about today would not be possible. We would also like to thank all the companies that work with us in our business including all the contractors and suppliers who provide the various services we require. These include equipment manufacturers, engineering companies, project management companies and suppliers large and small who cater to our every need as quickly as possible. Without all these partners we would not be as good as we are. Many of these companies are our sponsors this week. Thank you from all of us! This is also an appropriate time to pause and think about those Shaft Sinkers we have lost and those who have been injured on our projects around the world. Please bow your head and remember the Shaft Sinkers you have worked with, who did not return home or those who returned home injured. Think about how we can continue the work of making shafts safer every day as we go forward with our projects so everyone goes home to their families safe and sound. Pause for a minute and thank you. The subject matter we have chosen to discuss today is “Shaft Sinking in the Next Five Decades”. How will the work be done, how fast will it be done, what will it cost, and why we think this may occur? It is appropriate to consider this over the next few days with this auspicious group we have assembled. Today, we will look back at all the advances that have been made over the last many years and look ahead to the future to attempt to predict what shaft sinking will look like in 2070. When I look around the room I recognize that many of us will not be around then so we need not worry about being wrong. Suffice it to say that the next 50 years will be very interesting and exciting for the shaft sinking business."

Jan 1, 2019

By M. S. Diederichs, J. Oke

Surface limitations and a need for more direct transportation routes require serious consideration be given to the development of underground construction works in geotechnically unfavourable conditions. Due to geographical and logistical limitations, the construction of these underground works is restricted to difficult ground conditions that require temporary pre-support before excavation in order to avoid tunnel collapse or severe surface settlement. One type of pre-support is the umbrella arch system, which has been gaining popularity due to cost and time commitment in comparison with other pre-support systems (e.g., ground freezing). The umbrella arch system has been employed with significant frequency around the world, yet without a complete understanding about the mechanical response of the system. As such, literature has found the temporary support system to be both under and over designed with regard to safety and/or economical impact. The authors of this paper have made three major contributions to increase understanding of the umbrella arch system and to help optimize the design process. Firstly, the authors have developed nomenclature in order to categorize and define umbrella arch elements and systems, which also stimulated the creation of the empirically driven Umbrella Arch Selection Chart. Secondly, the authors have established a design methodology that is primarily based on advanced numerical modelling techniques to aid in the optimized design of the umbrella arch arrangement for site specific conditions. Thirdly, the authors have developed a semi-analytical model for umbrella arch systems employed in squeezing ground conditions by incorporating elastic foundation theory as well as convergence confinement theory. These three major contributions to such a widely employed tunnelling pre-support system will be introduced and explained in detail, and will include discussion of how each is incorporated in the complete tunnel support design process and throughout the construction and monitoring phase of a project.

Jan 1, 2015

By Ben Burger

This paper describes the design, the first three years of operation and optimization of a large, single stage SAG mill at Minera Yanacocha?s Gold Mill facility located in northern Peru. The Gold Mill was commissioned in March 2008 with a 32? x 32? EGL, 16.5MW single stage SAG mill, to process 5 million tonnes per annum high grade oxide Au/Ag ore. Operational control, feed size distribution, grinding media, mill liner, grate and pulp lifter optimisation have allowed the mill to operate 15-30% above design throughput rates, but with a product P80 coarser than design. As part of the optimization, full grinding circuit surveys were conducted. Historical operating data and survey data were analyzed and used to evaluate possible circuit changes and determine alternative operating strategies. A number of changes were implemented which allowed the SAG mill to operate similarly to a large ball mill and achieve finer mill product that was closer to design. This paper summarizes life to date data analysis and improvements made by Minera Yanacocha as well as the grinding circuit optimisation study conducted by Metso Process Technology and Innovation.

Jan 1, 2012

By A. J. Vivyurka

"This is a talk about the experiences that Rio Algom Mines Ltd. has had in the Elliot Lake uranium area trying to make the tailings more compatible with the surrounding environment. To get a better idea of the problem that we are associated with we have to go back to the location. The Elliot Lake uranium area is situated about half way between Sault Ste Marie and Sudbury and is about 20 miles north of Lake Huron. In 1950, uranium was discovered along the north shore of Lake Huron and by 1961 we had up to 9 mines operating in the area. Then the United States Atomic Energy Connnission withdrew their contracts so that at present there are only 2 mines operating. From all the mines that we had operating, we ended up with 7 tailings areas. From these 7 tailings areas we have 5 sites where we have active treatment or rehabilitation. There is the Pronto site where we are treating uranium tailings and copper tailings. There is the Lacnor-Nordic site which is located just out of Ellio Lake. The Milliken-Stanleigh site is in the Crotch Lake area. The Panel tailings is off the north end of Quirke Lake and finally there is the operating property of Rio Algom Mines at Quirke Lake.To understand fully the problem with this rehabilitation we have to look at the constituents of the ore. The ore body is a quartz conglomerate with uraninite and brannerite as the uranium bearing minerals in a fine ground mass of quartz, sericite and pyrite which separates the larger quartz pebbles. The approximate composition of the ore is: 80% quartz, 13% chlorite, 7% pyrite and 0.2% uranium minerals. The processing method also adds a few constituents to the final tailings."

Jan 1, 1975

By I. I. Betcherman

INTRODUCTION A DISTILLATION process offers one of the most effective methods of refining metals. Its use is particularly attractive in the case of reactive metals which are not readily treated by pyrometallurgical or electrolytic methods. The 1ow vapour pressure of most metals severely restricts the applicability of the disti1lation process and only one or two metals may be effectively treated under atmospheric pressure at a reasonable temperature. If operations are carried out under reduced pressure several additional metals may be effectively distilled. Table I shows the vapour pressure of a number of metals together with a value of the free energy of formation of their ?oxides. Zinc and cadmium head the list and for many years? they have .been effectively refined by distillation. Sodium, magnesium, and calcium follow in descending order, and they may be distilled readily in vacuo and with difficulty at atmospheric pressure. Sodium, which is not of great metallurgical interest, is not considered. Magnesium has recently been produced in considerable tonnages by distillation methods ( 3). This paper describes the vacuum distillation of magnesium and calcium, their separation from each other, and their separation from less volatile impurities.

Jan 1, 1951