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By Gheorghe Marin, Franco Chiesa, David Levasseur, Bernard Duchesne

Over the past 20 years, pressure die casting has evolved to a point where structural parts can now be produced by this process. This has been made possible by the development of vacuum die casting and by the development of die-castable, low iron content alloys. This resulted in the production of low porosity, weldable die-cast parts on which heat treatment may be applied. In the present paper, the influence porosities on the tensile properties of as-cast A360 tensile test bars was evaluated. This was done by casting 24 clusters, comprising three 6mm diameter test bars each, in a 280T cold chamber pressure die casting machine after a steady dynamic equilibrium had been reached. Modeling of filling allowed to explain the differences in porosity observed in the 3 samples of a same cluster while the results of solidification modeling were related to the microstructure. Tensile samples were radiographed on a real time X-cube apparatus and separated into 2 lots: samples where no porosity defect were visible and samples with visible defects. By measuring the tensile properties of the 2 lots it was possible to quantify the drop in tensile properties brought about by porosity, which is the most common defect observed in pressure die-cast parts.

Jan 1, 2015

By Tassos Grammatikopoulos, Sarah Prout, William Mercer, C. Gunning

"The Nechalacho rare metal deposit is located in NWT, Canada. Avalon Rare Metals Inc. is now evaluating processing techniques for recovery of REE to be used in the industry. QEMSCANTM analysis coupled with whole rock, REE, Zr and Nb, and electron microprobe analyses were utilized to identify and quantify the mineralogy of the deposit, and define the REE distribution among minerals. QEMSCANTM mineralogy was performed on fifty intact drill samples using EXPLOMINTM that provided textural information for the mineralization; crushed samples were used to determine the variability of the ore and define REE domains, and composite samples to determine the liberation and association of the REE minerals. The sample matrix predominantly consists of biotite, feldspars, quartz, muscovite/clays, minor chlorite and amphibole, magnetite and hematite and carbonates. REE include in decreasing order of abundance allanite (3.6%), monazite (1.5%), synchysite (0.9%), columbite (0.9%), fergusonite (0.6%), bastnaesite (0.4%), and zircon 11.0%. The majority of the HREE are carried by fergusonite and zircon, and the LREE by monazite, allanite, synchysite and bastnaesite. QEMSCANTM mineralogical data is extremely valuable in assessing the mineralogy of the ore. The data have been implemented in defining the mineralogy of the ore body, support metallurgical test work, predict recoveries and grades, and finally calculate reserves of the deposit.INTRODUCTIONAvalon is in the unique position of owning and developing one of the world’s largest rare earth deposits in the Northwest Territories of Canada.The term Rare Earth Elements (REE) is used to describe the 15 lanthanide elements or metals and, more recently, the metals yttrium and scandium, which display similar physical and chemical properties. REE are not particularly rare and have a similar abundance in the earth’s crust to nickel, tin or tungsten. Unlike these better known metals however, the rare earths have little tendency to become concentrated into commercial ore deposits. When they do become concentrated, deposits of REEs usually display enrichment in one of two groups. The most commonly occurring deposits of REEs contain a sub-group of elements referred to as the “light rare earth elements” (LREEs), which include lanthanum, cerium, praseodymium, neodymium, and samarium. Less commonly, deposits of REEs contain the much more valuable sub-group of elements referred to as the “heavy rare earth elements” (HREEs), which include europium, gadolinium, terbium, dysprosium, holmium, erbium, thulium, ytterbium and lutetium. Yttrium and scandium most commonly occur in association with this second HREE sub-group. Generally, LREEs are produced from the mineral bastnaesite, and the HREEs have been produced from the minerals xenotime and monazite. LREEs continue to be produced from bastnaesite, however in the 1980s, China started production of HREEs from unique laterite clay deposits. China now produces approximately 95% of the world’s rare earths."

Jan 1, 2011

By T. K. Dichmann, J. A. Finch

"Microflotation, laboratory batch, and in-plant minicell continuous tests were conducted to determine if copper ions not only activate sphalerite but also depress pyrite when the minerals are together. Microflotation tests showed that pyrite recovery increased with addition of copper ions in the pH range of 5-10 when alone, but decreased when sphalerite was also present. Batch flotation tests (Red Dog ore) showed that upon addition of copper, pyrite recovery decreased significantly. The continuous tests (Louvicourt Mine) confirmed this and indicated that with excess copper pyrite recovery can increase. The results suggest there is competition for copper and xanthate in favour of sphalerite. To take advantage of this competition it may be necessary to control the sequence and rate of copper, collector, and lime addition. INTRODUCTIONFlotation, in common with all physical separations, is an imperfect process that results in misplacement of minerals. One source of this imperfection is accidental activation by contaminating metal ions. An important practical example is the misplacement of pyrite to zinc concentrates. Flotation of sphalerite generally requires activation by copper ions, added as copper sulphate. Copper ions are also used to activate pyrite [l] hence there is the suspicion that accidental activation contributes to the recovery of pyrite (Py) with the sphalerite (Sp). However, recent literature has shown that copper ions may make Py less floatable if Sp is also present. Xu et al. [2] conducted continuous minicell tests in-plant at Kidd Creek Operations (division of Falconbridge Ltd.) and showed that while the addition of copper sulphate improved zinc/iron selectivity, it did so as much by reducing the recovery of pyrite as increasing zinc recovery (Figure 1)."

Jan 1, 2000

By Jon R. Maki

"EVTAC, an iron ore mine in Minnesota, had trouble meeting the customer’s pellet silica requirements. In addition, the mine plan was moving towards a lower grade ore and funding was not available for plant improvements to compensate for the change in ore or improving quality. Small changes to optimize various processing equipment resulted in significant improvements in both quality and productivity; for example, productivity on one mill line increased by 10% while producing lower silica concentrate. This paper will provide the methodology that allowed EVTAC to improve quality and productivity through process optimization with little capital investment.INTRODUCTION EVTAC, an iron ore mine in Minnesota, had trouble meeting the customer pellet silica requirements. In addition, the short-term mine plan was changing to a lower ore grade. Because of limited funds, EVTAC did not have the capital for investing in the plant to compensate for the ore change or to make improvements to lower product silica levels. Small changes to optimize process equipment resulted in significant improvements in quality and productivity; for example, one mill line feed rates increased by 10% while also producing a higher quality concentrate. These changes were sufficient enough to allow the mine to lower the target ore grade beyond their initial goal. At one time, a $3M flotation facility was considered as a way to meet quality with the new ore grade. However, the process changes were able to accomplish the same silica reduction with roughly a $50,000 investment in VFDs for sump controls. This paper will discuss some of the modifications that allowed EVTAC to meet the quality and production objectives with little capital investment. It also discusses the methodology used to make these process improvements."

Jan 1, 2014

By C. V. Umipig, G. G. Hutalle, E. E. Israel

TVI Resource Development (Phils) Inc. Canatuan mine is a polymetallic ore deposit located in the Province of Zamboanga del Norte on the island of Mindanao in the Philippines. The deposit consisted of an oxidized capping of gossan overlying a Besshi type volcanogenic massive sulphide lens which is near surface and flat lying. Construction and development of a sulphide processing plant began in early 2007 and commissioned in November 2008. During the initial stages of sulphide operation, high copper (Cu)-low zinc (Zn) ore was fed to the plant. Marketable copper concentrate at 91-95% recovery was produced utilizing a high-pH bulk sulphide rougher flotation and 2 cleaning stages. Because of increased levels of Zn of the ore body, a Cu/Zn separation circuit was initiated in early 2010 following an intensive laboratory test work program. This Cu/Zn separation circuit was necessitated because of the copper ion pre-activation of the sphalerite that lead to the zinc being readily recovered to the copper rougher concentrate. In-situ activation of zinc (in the pit and stockpiles) was caused by the leaching of secondary copper minerals in the ore (particularly from the dominant copper species, chalcocite and covellite). This problem required the development of metallurgically complex and intensive reagent regime to effectively depress the activated sphalerite. As well, the immediate reactivation of zinc after it has been depressed, at the onset of Cu/Zn separation flotation, added additional metallurgical complexity to the process. Subsequent to the Cu/Zn separation circuit, an unconventional, yet proven effective, simultaneous cyanide detoxification-zinc flotation circuit was set-up to recover the zinc. Laboratory test work results realized Cu-Zn recovery differential from a Cu/Zn separation circuit of over 40-45%. This performance has been achieved in the plant, at 72% Cu overall recovery. With the gradual improvements on-going with zinc circuit capacity and effective zinc separation being achieved, zinc recovery has currently reached 40%.

Jan 1, 2012

By Ian Orford, Peter Radziszewski, Lud Strah

"An autogenous grinding circuit was commissioned at the Brunswick concentrator in October 1998. The AG mill was subsequently converted to SAG operation in May 1999 for a number of reasons. The mill had 48 rows of the NORCAST top-hat integral lifter/liners configured in Hi-Lo mode. A charge motion simulation study, the results of which are the subject of this presentation, was initiated to develop a better understanding of SAG mill behaviour. Using a design of experiments approach, a lifter face angle was determined that would increase mill power draw. Once determined, lifter profiles were modified and installed in July 2001, and a second set installed in February 2002. Resulting changes in mill power consumption are presented.INTRODUCTIONThe Noranda Inc. Brunswick Mine, located near Bathurst, New Brunswick, is a large lead/zinc mine and began milling ore in 1964. The ore is massive sulphide with approximately 20% valuable minerals, 60% pyrite and 20% silicates. In October 1998, an autogenous grinding circuit was commissioned at the Brunswick concentrator (Cooper, Orford, Larsen, 2001).At the time, the autogenous grinding circuit replaced existing surface crushing plants, fine ore storage, rod mills and primary ball mills. The 8.54 m diameter x 4.27 m long (EGL) ANI mill was subsequently converted to SAG operation in May 1999.The mill had 48 rows of the NORCAST top-hat integral lifter/liners and was operated initially with all shell lifters at the same height. Following problems with packing and abnormal wear pattern, the profile was changed to a high-low configuration.Further, it was found that over the life of the lifters, mill power draw and throughput varied considerably. Although a change in mill power draw was expected following lifter replacement (predicted during mill design to be in the order of 300-400 kW), the replacement of the worn lifters with new high lifters resulted in mill power draw decreased by some 1000 kW."

Jan 1, 2003

By G G. Suffel

T HESE remarks on the occurrence of strata-bound massive sulphide deposits are submitted be-cause the writer recently visited the principal Cyprus deposits (in January, 1965) and has also a back-ground of many years spent largely in detailed observation and map-ping of the Horne and Amulet ore-bodies, now largely inaccessible. They are particularly related to the Symposium papers (September Bulletin) by Hutchinson and by Roscoe. Some comments are offered on the cupriferous pyrite deposits of Cyprus. As the writer is currently reviewing the published literature on the related deposits in Turkey, a brief account of some characteristic Anatolian sulphide occurrences is given. The discussion is closed with some thoughts on the relationship between these Mediterranean sulphide deposits and those of northwestern Quebec and the Superior province in general. The geology of Cyprus, as it relates to the sulphide deposits, is well summarized by Hutchinson (1965). After visiting Mavrovouni, Skouriotissa and Limni and after much thought on the possible ge-nesis of these ores, it seems quite logical to me that he is right. They are indeed closely related to late-stage submarine volcanism.

Jan 1, 1965

By Niklas Frank

Ore and waste transports form an integral part of any mining operation and represents a substantial portion of the mine´s total operating cost. Mine layout, truck fleet and size, production rates, haulage distances, shift schedules, maintenance strategies etc. are all important and not at all static variables. They all change during the life of a mine and interact in a dynamic way. It is therefore essential for the successful mine management to constantly analyse and optimise the mine's transport systems. Önnerlöv has recently conducted a life-of-mine haulage simulation project for Zinkgruvan Mining, a medium sized, underground zinc producer, located approximately 60 kilometers south of Örebro, Sweden. The main objective of the project was to analyse and optimise the investments required to provide ore and waste haulage under changing conditions over an +10- years period: Deepening of the mine, increased production rates, changed mine layout, changed truck fleet etc. A large number of scenarios has been tested in the simulation model in order to identify the "lowest cost per ton" solution. This paper comprises an introduction of Zinkgruvan Mining and a brief background to the actual project, a description of the simulation model, project conclusions and some useful hands-on tips for mining engineers and project managers how to run a successful simulation project.

May 1, 2003

By Ilpo O. Auranen

Background: Mines of today are moving ever increasing tonnages of rock. The new ore bodies tend be of lower grade and located deeper in the earth?s crust. The industry?s focus to overcome this ?deeper but poorer? ore problem has traditionally been to lower the costs by increasing the scale of mining. This has resulted in an ever increasing size of mining equipment like drills, loading shovels, dump trucks. The main focus has been in reducing the cost per ton of mining. This cost focus has often resulted in increased waste rock dilution and more ore losses. From costs to value: Waste rock dilution means waste rock in the concentrator feed. It?s value in open pit mining is equal to the value of ore it replaces. In underground mining, waste rock dilution?s value is equal to the value of ore it replaces added to the costs of mining it. It is commonly accepted that waste rock dilution in open pit mining ranges between 5 and 20% and in underground mining between 5 and 40 % . In a typical large open pit mine ore value is from 20 to 60 USD per ton and mining costs are 1 to 3 USD per ton. One percentage point reduction in waste rock dilution represents savings corresponding to about 20% of total mining operation costs. The difficulty of reducing waste rock dilution comes from the complicated nature of the ores. Accurate ore waste-border detection has not been possible due to excessive costs. Core drilling combined with geochemical lab analysis is the most accurate method for locating ore waste borders. However, it is also expensive and time consuming, which is why it is often combined with somewhat lower cost methods like sampling drill cuttings from blasting drill holes or from reverse circulation drills and analyzing them in a laboratory. The main drawback with these methods, in addition high costs, lies in the sampling accuracy and in the quality control of the whole process when a large number of samples have to be handled manually. In every mine it is possible to halve waste rock dilution by using new on-line analysis technology .

Oct 1, 2009

By R. J. Balfour, T. L. Papucciyan

"INTRODUCTIONCapital Cost considerations are basic to feasibility in the achievement of success for Mineral Processing Projects as much as or more than in many other industries. Capital cost is an important factor ae it 1R the unit of measurement for process engineering economic decisions. As a result of ever increasing complexity, alternatives of process and remoteness of site, there is an ever increasing emphasis on accurate estimating and the attendant cost control systems in the international market of today for the Mineral Industry where initial plant costs are becoming a more important factor than operating costs. An undervaluation means trouble ahead. An over-cautious assessment may bring an unnecessary abandonment of a project. Consequently, the ultimate cost must be predicted early and accurately and must be controlled to this parameter during the design and construction phases. Cost Est!~ mating is one of a variety of project management tools used to assist in planning and achieving the optimum balance between Quality, Cost and Time, the eternal triangle of evaluation, namely obtaining a project of acceptable quality at minimum cost and on time. In order to attain these established objectives, proper planning, organizing, coordinating, monitoring, adjusting and controlling the design, procurement and construction elements of the project are essential.This project management approach based upon the principle of a Task Force under single management, comprising a complete service organization, requires a substantial number of qualified individuals specialized in their own field and capable of working together as a team to perform their various functions effectively. The Task Force is supported by their chiefs of discipline and performs in accordance with established procedures and standards under the leadership of a Project Manager.General Engineering Company Limited with the backing of the full resources of Surveyer, Nenniger & Chenevert and its total staff of approximately 1,0009 offers the complete project management service to the Mineral Processing Industry of North America and the rest of the world. As mine sites become more remote and difficult of access, the need for an experienced team of specialists who work under the Task Force principle, with well established standards and procedures becomes even more important. Our long history of successful pro1ecta in this field has been our training ground for the fulfillment of this need."

Jan 1, 1972

By R. B. Throndycraft

"1. Belt-Drop Agglomeration Preferred at Heap Leach Operations. The agglomeration method now preferred at heap leach operations utilizes belt-drops along conveyors. Belt-drop agglomeration is being designed into numerous upcoming heap leach operations. It offers substantial capital cost savings over previously favored drum agglomerators. It has such significant operating and maintenance advantages that some drum agglomerators have been retired in favor of belt-drops.The majority of heap leach operations are now built with crushing included in the flowsheet. If crushing is included, belt-drop agglomeration can be added for almost no cost. The agglomeration equipment is simply a few conveyor transfers, designed to provide the tumbling and rolling action which forms the agglomerates. Controlled water sprays are needed at each of these agglomeration transfers. Cement or lime is added as a binder, to stabilize the agglomerates. This reagent is also needed to raise pH to safe operating levels.Addition of the pH control reagent before stacking the ore offers a reagent cost benefit. In the belt-drop mixing technique, the ore is intimately mixed with the lime (or cement). After stacking on the heaps, when the cyanide solution is sprayed onto the high-ph ore, very little cyanide is consumed. Without the pre-addition of lime or cement, the leach solution is sprayed onto neutral ore. The mixture of high-pH leach solution and neutral pH ore causes the solution pH to drop, releasing cyanide gas. This release is minor from a safety viewpoint, much less than dangerous, but the gas release is a costly waste of the cyanide reagent. Pilot-scale comparison tests were made at Pinson. Ore pre-treated with lime or cement showed 30% less cyanide consumption than untreated ore."

Jan 1, 1989

By E. H. Henderson

Introduction Stamp-Milling and amalgamation is, perhaps, one of the oldest processes of recovering gold from its ores. Many papers, theses, and textbooks have been written on the subject of stamp-milling practice in general, but, to my knowledge, there have been no publications describing the modern stamp-mill as designed, developed, and operated in Nova Scotia. Although it may be said that the gravity stamp-mill is the most primitive, and perhaps the crudest, crushing machine in existence, it is at the same time the most satisfactory all-round milling and amalgamating device yet contrived. Originating in Saxony, and used in North America first in Virginia and Georgia, the stamp-mill, in a primitive form, later found its way to California and Colorado. Here, it speedily developed from an affair of wood shod with iron, and operated with the crudest gearing, to substantially its present form. Briefly, it consists (see Figure 1) essentially of a series of hammers (the stamps), which fall upon a mixture of ore and water contained in a cast-iron box (the mortar). The resulting crushed ore (the pulp) is splashed out through a screen on the front of the box. For some years after the discovery of gold in Nova Scotia, in 1861, the ? arrastra and Chilean mill were used to recover the gold from the rich ores. The stamp-mill followed, and various types were used, principally the California and Gilpin County types. Then, in 1895, Mr. M. R. O'Shaughnessy, one of our foremost amalgamators, conceived the idea of a new type of stampmill, which has since become standard in Nova Scotia gold mines. This mill, unique in its design and operation, is specially adapted for the most efficient extraction and recovery of gold as it occurs in our Nova Scotian ores. It is commonly known as the Underwater Feed type, as opposed to the usual Overwater or Top Feed type, and it may be rightly termed The Nova Scotia Self-Classifying Stamp Amalgamation Mill. It is conceded by experienced mill-men that it is superior to any other type of mill for recovering the very coarse and granular gold typical of the ores of this Province. The particular installation and practice to be described in this paper is that of the Guysborough Mines, Limited, at Goldenville, Guysborough county. Emphasis will be laid on certain operating innovations of extreme simplicity and high efficiency.

Jan 1, 1935

By Paul Gauthier

Aujourd?hui, avec un faible prix de l?or sur le marché, les opérateurs miniers doivent revoir leur façon de faire. La méthode de minage «Eureka» permet de réduire les coûts de production et d?accélérer la mise en production des chantiers. Les principales étapes du cycle de minage et les problématiques associées à l?utilisation de cette méthode à la Mine Bousquet y sont décrites.

Oct 1, 2001

By O. Rabeau, M. Legault, A. Cheilletz, L. Z. Cheng, J. J. Royer, M. Jébrak

"By compiling geological, structural, geophysical, and geochemical information into a 3D geological model, we evaluated the orogenic gold potential in the vicinity of a hidden segment of an important Archean fault zone, the Cadillac-Larder Lake fault (CLLF) in the region of Rouyn-Noranda. The segment of CLLF in the present study is partly covered by Proterozoic sedimentary rocks. Because more than 2000 t Au have been extracted along the CLLF to date, our objective is to evaluate the gold potential at depth along a poorly known segment of this fault. A 3D geological model (50 km × 9 km × 1.5 km) including the covered segment was built through the compilation and homogenization of available geological data and the construction of 23 cross sections. The geology under the Proterozoic cover was evaluated using geophysical inversions, drill holes (42 in total), and surrounding geology. All available assays were filtered and upscaled to a 250 m × 250 m × 250 m regular cell grid to determine and quantify spatial relationships between geological features and mineralized occurrences using the weights of evidence method. Structural features, such as E-W-trending faults and fault intersections, and certain lithologies with a high primary porosity such as volcanoclastic rocks of the Blake River Group and Timiskaming sedimentary rocks, proved to be very prospective, yielding favourable factors with a weight of evidence index W+ > 0.24. These salient features were then assigned a combination index for ultimately evaluating the orogenic gold potential under the sedimentary cover. The zones resulting in an optimization of exploration targeting were attributed the highest probability, representing ~1% of the initial volume. La compilation des données géologiques, géophysiques, structurales, et géochimiques et leur intégration au sein d’un modèle géologique 3D a permis d’évaluer le potentiel aurifère dans les environs d’une importante zone de faille archéenne sous couverture protérozoïque. Cette méthodologie a été développée en utilisant le cas de la Faille Cadillac-Larder Lake (FCLL) dans la région de Rouyn-Noranda. Puisque plus de 2000 t d’Or ont été extrait à proximité de la FCLL, l’objectifs des travaux est l’évaluation du potentiel aurifère en profondeur de ce segment. Un modèle géologique 3D (50 km × 9 km × 1.5 km) incluant le segment sous couverture de la FCLL a été construit en utilisant l’ensemble des données géologiques disponibles et 23 coupes géologiques. La géologie du socle archéen sous la couverture protérozoïque a été interprétée avec des inversions géophysiques, des forages (42 au total) et les contacts géologiques exposés. Les teneurs aurifères dans le secteur ont été filtrées et ajusté à une grille régulière de 250 m × 250 m × 250 m afin d’évaluer et de quantifier les relations spatiales existantes entre certaines entités géologiques et les emplacements minéralisés en utilisant la méthode du poids de la preuve. Les failles de direction Est-Ouest, les intersections de failles ainsi que certaines lithologies possédant une porosité primaire importantes comme les certains tuf du Groupe de Blake River et les sédiments du Timiskaming se sont montrés très prospectifs avec des valeurs de W+ > 0.24. Ces entités géologiques ont ensuite été utilisées par le biais d’un indice de combinaison pour évaluer le potentiel en or de type orogénique sous le couvert sédimentaire. Cette méthodologie permet une optimisation des campagnes d’exploration puisque les zones pour lesquelles la probabilité de minéralisations a été évaluée à la plus haute valeur représentent 1% du volume initial."

Jan 1, 2009

By L. Su, S. Liu, K. Jiang, K. Xie

The reduction of iron dissolution and acid consumption is fairly important in the extraction of Ni and Co from laterite ores from both economic and efficiency perspectives. BGRIMM conducted detailed research to investigate the behavior of iron in a nickel laterite leaching process. The mineralogy of Ni and Co was studied. In limonite ore about 76% of nickel is present in goethite while in saprolite ore 79% of nickel is associated with magnesium and silicon minerals. Test work revealed that limonite can be decomposed completely with enough acid concentration during atmospheric leaching. More than 98% of Ni and Co were leached into solution while 85% of the iron was also dissolved with 130-140 g/L iron in solution. Iron precipitates as hematite in the temperature range 250-270?C without neutralization reagents. That’s why limonite is treated by high pressure acid leaching (HPAL) technology to achieve nickel extraction and iron precipitation in the same autoclave. BGRIMM’s test results indicated that the remaining acid in atmospheric leaching solution and acid released by iron precipitation at a lower temperature was enough to leach saprolite efficiently. The so-called inverse leaching technology leaches limonite at atmospheric condition and the leaching slurry combined with saprolite is pumped into an autoclave to leach saprolite at 150-170?C. In the autoclave iron precipitates as hematite without acid consumption. Total nickel and cobalt extractions in the test program were more than 90% with iron content in solution below 8 g/L. In this inverse process iron dissolves in the atmospheric leaching stage and precipitates as hematite in the low temperature leaching stage. Compared with HPAL the leaching temperature is reduced to 150-170 ºC from 250-270 ºC. Compared to atmospheric tank leaching acid consumption is reduced from 850-950 kg/t ore (atmospheric) to 550-650 kg/t ore (inverse).

Jan 1, 2016

By Johnny Zhang, Yang Anguo, Li Weifeng, Xueyi Guo

"SKSTM oxygen bottom blowing smelting technology was developed by SKS-ENFI at SKS Smelter, Hunan in the 1990s to replace the conventional sintering process and thus to meet the high national EHS standard for lead.Since then, two major milestones have been achieved:1. Since 2002, the SKS oxygen bottom blowing smelting-blast furnace reduction has been in operation at SKS, YGL, Chizhou smelter etc. Now in China, about 50% of the lead (4.2 millions t/year) is produced using SKS process; and2. Since 2008, the blast furnace has been replaced by direct reduction of molten rich lead slag at YGJ smelter, which is more energy-efficient since no slag cooling is required. This two stage process (SKS Oxidation- Direct Reduction) is named as the YGL process.Since 2008, numerous technological innovations have been implemented in the furnace design, process control, and maintenance at YGL. The plant has been successfully demonstrated its unique excellent performances on all aspects from processing capacity, product quality, energy reduction, cost reduction and emission reduction. The purpose of this report is to update all the improvements with proven tracks.Meanwhile, a preliminary comparison on the following aspects between the YGL process and other similar technologies has been discussed:1. Plant design and operating conditions;2. Mass and heat balances, metal (Pb, Ag) recovery;3. Consumables, oxygen gun and refractory lining life;4. Energy consumption and operating costs;5. Sulfur capture and emission reduction etc.It is concluded that the modified SKSTM process is a clean SO2-free emission process. The process is characteristic of other features such as no extra fuel addition, low capital with a small foot-print design, low OPEX, high throughput, easy operation, long refractory lining life and high productivity (95%+ on line time annually). It offers an alternative cost-effective technology for lead smelting industry."

Jan 1, 2014

By Nehad Elsafty

It is always necessary to expand existing buildings, so that they could meet the growing societal needs. Compared with demolishing them and then constructing new ones, the expansion of existing buildings could not only reduce project cost, but also save the use of energy and raw materials in projects. Therefore, it is widely adopted in most under-developed countries. However, the expansion of existing buildings is not an easy task. It has been involved with a lot of issues in defining the scope of work, planning landscape, managing construction activities, etc. This paper investigates the potential challenges in expanding existing buildings. The project of expanding the Al-Agoza Hospital, one of the biggest and oldest hospitals in Cairo, Egypt, is selected as a case study. In the case study, the project documents have been carefully reviewed. The project engineers and the health-care professionals working in the hospital have been interviewed. This way, the point of views from the owner, the architect/engineer, and the contractor have been taken into account. Several lessons are learned through the discussion with these people. The results from this paper are expected to provide pieces of useful advice and suggestion for the expansion of existing buildings in the future.

Aug 1, 2013

By Scott Martin

This paper will discuss the challenges currently facing the mining industry. By sharing experiences gained in the CIM Leading in Mining Course, I will highlight what’s working, what’s not, what are the challenges, what are the concerns, what are the opportunities, what needs to happen. Specifically, I will discuss what is currently happening with leadership and leadership development in members’ own organizations, what skill sets are important to develop in different hierarchical groups and how comprehensive leadership courses like the leading in Mining course can help improve leadership within individuals and the industry. Leadership development is a process not an event.

Jan 1, 2014

By Dave Marsh, John Goode, Niels Verbaan, Kevin Bradley, Henrie Notzl

The mineralogy of the Avalon Nechalacho deposit is complex with significant fractions of HREEs, Zr, and Nb found in refractory minerals such as zircon and fergusonite. Sulphuric acid baking followed by water leaching can be used to extract REEs from flotation concentrates, but acid baking does not completely "crack" the zircon and fergusonite, so the extractions of Zr, Nb, and HREEs are not high. An alternative flowsheet has been developed that is able to obtain significantly improved extractions of HREEs, Zr, and Nb from flotation concentrates or acid bake residues. This is accomplished by cracking the zircon and fergusonite minerals with caustic soda at high temperatures, followed by water washing to recover excess caustic soda and soluble sodium silicates and phosphates. REEs, Zr, and Nb in the solid residue can then be solubilized with HCl or H2SO4. This paper highlights some of the findings from an extensive laboratory program and discusses several flowsheet implications.

Jan 1, 2014

By A. M. Ramadan

The composition of phosphate rocks, mainly used in manufacturing fertilizers, varies among deposits; therefore, phosphate rocks from different sources are expected to behave differently in acidulation processes. This paper examines the effect of preconcentration of phosphate ore from El-Sebaiya East, Egypt, using flotation on uranium leaching. A high-grade ore sample (27.3% P2O5, 168 mg/kg U) supplied by the Al-Nasr Mining Company, was subjected to reverse single-stage flotation to separate carbonate from the phosphate sample. Subsequently, P2O5 and U concentrations were 32.2% (92.1% recovery) and 225 mg/kg, respectively. After the leach process, the U concentration in the liquor was 319 mg/L.

Jan 1, 2014