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By Q. Han, W. Dai

Al-Ti-B type master alloys have been widely used in the grain refinement of aluminum since 1940s. The introduction of Al3Ti and TiB2(AlB2) particles reduces the grain sizes down to about 200 micrometer level and makes cold work possible. However, the mechanism for the grain refinement is still not clear, though it is believed that TiB2 in the presence of Al3Ti nucleate a-Al grains in solidification. This paper presents our finding on the formation of (Ti,Al)B2 solid solution, which leads to a proposed theory on grain refinement by Al-Ti-B type master alloys that upon addition into aluminum melt stable TiB2 particles react with aluminum slowly and release titanium into the melt. The titanium thus released in combination with titanium in the melt through dissolution of Al3Ti maintains a transient titanium-rich layer on the surface of (Ti,Al)B2. This layer offers a low crystal mismatch with a- Al and promotes the nucleation of aluminum grains.

Jan 1, 2014

By G. Ladouceur

"Northgate Patino Mines produces a copper concentrate assaying 22% Cu. Moisture content of the product averages 12.5% H20 when dewatering with conventional vacuum disk filters. Recent smelter contracts reflect higher energy costs and increased moisutre penalty charges. This situation forced Northgate to investigate possible alternatives such as rotary dryers and pressure filtration. After laboratory testing, it was found that pressure filtration could economically meet the smelter's requirements. Pilot plant testing was then undertaken and confirmed the laboratory test results. Based on these findings, a model I.S.F. 1500 x 28/22 I.R. Lasta filter press was purchased.This paper describes the Copper Rand dewatering flowsheet the operation of the press and the economics of the process.IntroductionNorthgate Patino Inc. operates 2 und erground copper mines and a 3300 ton per day concentrator located in Chibougamau, Quebec, some 425 miles north of Montreal and 300 miles northeast of Noranda, Quebec.Copper concentrate is produced at a rate of 250 tons per day and trucked to a C.N. siding 10 miles from the mill site and then railed to the Noranda Smelter.Until early 1981, moisture content of the concentrate at 12.5070 was acceptable since smelter penalty charges were relatively low ($0.10 per ton per unit over 10070 moisutre). In January 1981, a new contract was signed in which the penalty was raised to $1.10 per ton per unit over 8070 . This and the increasing truck and rail costs justified the change-over from vacuum to pressure filtration . The Lasta I.S .F. 1500 x 28/22 was then purchased and installed, and operation of the plant started in December 1981."

Jan 1, 1987

By Jean-Luc Beaudoin

"In addition to the actual energy consumed, billing is dependent on the peak load and the power factor. The lower the power factor goes, the higher will be the penalty the consumer pays. The principles of a power factor regulator, the two ways of improving the power factor and methods for calculating the capacitor values required will be described.The rate structure of the utility company is founded on three cornerstones: the energy consumed during a given period of time and two penalties imposed when the power factor is low or the power demand is high. Even though the normal consumption of a customer might be 10 kW, with a peak demand of 20 kW only once a month, the wiring and associated equipment at the input to the customer's plant will have to accommodate this peak demand. The power company has no means of forecasting these demands and cannot therefore provide the additional equipment necessary. So it exacts a penalty, and the customer must take all possible steps to avoid these peak power demands. This can be accomplished by using an automatic peak load regulator which will selectively disconnect the loads which are not essential in the operation of a plant .The principles of such a controller will be described.Principles of power factor regulatorsThe operating principles of a power factor regulator will now be described. It might be asked why the power factor of an electrical installation in a mine, or a factory, should be corrected at all, but there is a simple answer. In addition to the actual energy consumed, the billing is dependent on the peak load and the power factor. The lower the power factor goes, the higher will be the penalty the consumer pays."

Jan 1, 1982

By Godfrey W. McDonald, Dennis E. Fredrickson

"Commercial coal mining has been recorded in Nova Scotia since 1863 and during this period an estimated 60,000,000 short tons of mine discard containing 10 1030% coal has been accumulated, Approximately 70% of these dumps are in the Cape Breton region and most are owned by the Cape Breton Development Corporation (CBDC), formerly DEVCO.The explosion of oil produced energy costs which created an increase in thermal coal demand as well as the desire to rehabilitate abandoned mine dumps encouraged coal reclamation. CBDC was receptive to a proposal from the Selminco Joint Venture and encouraged the building of the Princess Recovery Plant (1980) and The Summit Plant (1982).The mine dump contains scrap, waste rocks, coal and substantial quantities of fine shale. The scrap and oversize are removed and the remaining material slurried with water to a density of 1.45. The slurry is cycloned to separate the saleable thermal coal from the rejects which are returned to designated areas of the dump for contouring and revegetation.BackgroundThe recorded coal production in Nova Scotia from 1863 to 1980 was 390,000,000 short tons. This extraction was associated with an estimated 60,000,000 short tons of discarded material. This mining discard and coal washing plant reject are located in numerous, varying sized dumps throughout the mining areas. Approximately 70% of these dumps are in the Sydney coal field region of Nova Scotia. The coal content of these dumps varies from 10 to 30% coal with an average of 18%.Energy CostsThe explosion o f energy costs in the mid-1970s greatly increased the sale price of thermal coal which made recovery of coal from some of the dumps economical. The environmental benefit of these operations will be the rehabilitation of the dumps upon cessation of the coal reclamation operation."

Jan 1, 1983

By John Clancy, Andrew J. Freyman

"Trends in world coal markets indicate that the next twenty years could be a period of intensive development for British Columbia's coal resource. This paper examines demand projections for metallurgical, thermal and chemical feedstock coals, with particular reference to domestic and Pacific Rim markets. Supply possibilities, from the point of view of British Columbia's competitive position, are explored, with emphasis on the interaction of such factors as capital, labour, port and transportation requirements.The paper concludes with an examination of British Columbia's strategies in light of the foregoing, and comments on the government's role in ensuring that development of the coal resource be attained in a manner consistent with over-all provincial goals and objectives.IntroductionFor the purposes of this paper the Pacific Rim is defined as including, in addition to Canada, the following: Japan, Korea, the People 's Republic of China, Hong Kong, Taiwan, the countries of the Indonesian Archipelago, Australia, New Zealand and the Pacific-bordering countries of South America and Mexico. The United States has been excluded in this definition in order to better illustrate the economic growth potential for British Columbia coal in the Pacific Rim economies.In the past decade, the Pacific Rim economies have grown faster than most other economies in the rest of the world; economic growth rates of 10 per cent were common in the region prior to the 1974 slowdown and the average rate of economic expansion in the mid-1970s has exceeded 6 per cent. Rapid expansion of international trade in the region has been a major stimulus in this process. By 1979, Pacific Rim two-way trade has reached an estimated $550 billion-about 20 per cent of world trade."

Jan 1, 1982

By D. H. Whiting, A. J. Sinclair

"The San Antonio gold mine, in the Archean Rice Lake Greenstone Belt, was in continuous production from 1932 to 1968, during which period the mine yielded 1.35 million ounces (41 985 kg) of gold and 195,000 ounces (6065 kg) of silver from 4.88 million tons (4.44 Mt) of ore for average grades of 0.28 ounces per ton (9.0 g/t) gold and 0.04 ounces per ton (1.3 g/t) silver. After the 1968 closure, several publications appeared implying that gold grade of ore decreased with increasing depth.Recent observations by the authors show that the relatively low grades in the middle levels of the mine are controlled geologically. However, the reduced average are grade in the lower levels of the mine was due to inadequate mining control that led to a high level of dilution. The similarity of geological factors such as vein styles within structural domains, variations in mining control, recent ore muck sampling results and lithogeo chemical profiles all indicate the similarity of ore in both upper and lower levels.Strict mining control should yield improved grades of gold for future production from the lower levels at San Antonio, in contrast with the production grades of the late 19605that represented extensively diluted ore. High-quality production records are an essential component in re-evaluating a past producer, particularly where workings are accessible. Poorly founded conclusions which cannot be checked by access to workings can have a negative impact on the evaluation procedure."

Jan 1, 1990

By D. C. Brady

"Within six months from award of a project management contract, Les Mines Dumagami Limitée, Cadillac, Quebec, was operating its new hoisting facility. This was accomplished with less than a six-week shutdown of mine production while still maintaining shaft access for underground maintenance.The work involved installing a refurbished /3.5 ft diameter double drum hoist, replacing a timber headframe with a newsteel headframe, increasing the ore bin capacity from 500tons to 1000 tons, constructing a new substation, and upgrading related facilities.This paper describes the work and methods used by various contractors, Les Mines Dumagami Limitée and J.S. Redpath Mining Consultants Limited to bring this project to a successful completion. It discusses the history of the project from proposal stage through to the design, procurement, scheduling, and construction management. The paper also comments on problems that were encountered and the methods used to overcome them.IntroductionThe Donald J. Laronde Mine of Les Mines Dumagami Limitée (Dumagarni), a gold and copper producer, is located in northwestern Quebec, approximately half way between Val d'Or and Rouyn. The mine started production in the spring of 1988 and the first gold bar was poured in June 1988.Following a request by Dumagami, on February 22, 1989, J.S. Redpath Mining Consultants Limited(Redpath) delivered a proposal to increase hoisting capabilities from 1500 tpd to a minimum of 2500tpd. Although various concepts were considered, Redpath proposed a new hoist house along the same shaft axis as the existing hoist, and located on the site of the existing substation. They would then provide a new substation, replace the headframe with a largersteel structure , gain extra storage capacities by increasing the bin heights, and raising the roof of the bin house accordingly.Although underground mining would have to be stopped for a month, the mill production would be maintained by feed from Dumagami's open pit.Figure 1 shows the original plant layout and Figure 2 represents the general arrangement after the project was completed . Table 1 compares the original hoisting facilities to the new facilities.On March 2, Dumagami awarded Redpath a contract that included preliminary engineering, detailed engineering, procurement and contract administration, and later added construction management. The preliminary engineering refined the initialconcept, selected major equipment, and prepared schedules and capital cost estimates for the total project."

Jan 1, 1991

By Jean-Luc Collins

"The paper deals with the impact of fleet imbalance on a deterministic evaluation of shovel fleet productive capacity. It shows that if the fleet imbalance is overlooked, then estimated fleet productive capacity will be overrated.The paper shows that fleet imbalance inflates utilization, and care must be taken when using such statistics which are derived from reported field-equipment-hours. It shows that the utilization must be deflated by a factor called 'balancing efficiency"" in order to estimate shovel productive capacity correctly. The paper also shows how fleet imbalance can be controlled by managing idle time or downtime in an actual operation.The technique used for measuring the balancing efficiency factor is based on statistical analysis of equipment utilization."

Jan 1, 1980

By Lluís Fontboté, François Lieben, Robert Moritz, Massimo Chiaradia

"Abstract - The Las Cañas stratabound Pb-Zn deposit, in northern Chile, is hosted by the Lower Cretaceous back-arc marine carbonate and subordinate andesitic volcanic and volcaniclastic rocks of the Chañarcillo Group, one hundred meters above a thick pile of magmatic arc-related basaltic to andesitic lavas of the Bandurrias Formation. The deposit is not obviously related to local intrusive rocks, which include an Upper Cretaceous granodiorite batholith and dacite dikes emplaced along normal faults. Stratabound Pb-Zn-Ba mineralization is constrained within several porous stratigraphic units, including a 50 cm thick lapilli-tuff and less well mineralized algal-laminated limestone. Veins that locally carry some Pb-Zn occur along normal faults that displace the stratabound orebodies, and this study shows that they represent a different type of mineralization.Fluid inclusion microthermometric data for ore-stage quartz and barite indicate saline fluids (6 to 30 wt% NaCleq) and a formation temperature in the range of 80°C to 160°C for the stratabound mineralization. Two types of fluid inclusions occur in the veins. The type I vein fluid inclusions indicate less saline (average 7 wt% NaCleq) and slightly higher temperature fluids, in the range of 110°C to 220°C. The type II vein fluid inclusions indicate highly saline fluids (25 to 40 wt% NaCleq) which might be related to minor copper mineralization.206Pb/204Pb, 207Pb/204Pb and 208Pb/204Pb ratios for galenas from the stratabound mineralization, with mean values of 18.42, 15.59, and 38.25, respectively, fall within the range of stratabound polymetallic deposits of northern and central Chile, and indicate a Lower Cretaceous magmatic rock source. Galenas from veins have slightly more radiogenic 206Pb/204Pb, 207Pb/204Pb and 208Pb/204Pb ratios, with mean values of 18.47, 15.61, and 38.34, suggesting two distinct episodes of mineralization.?34S values for galena and barite range from negative values (-31.2‰ to -12.5‰ and -7.7‰ to +12‰, respectively) in the mineralized tuff to higher values in the algal-laminated limestone (-2.5‰ to +1.2‰ and +13.2‰ to +17.4‰, respectively), close to Lower Cretaceous marine sulfate values for barite, and homogeneous values for the galena from veins (-12.0‰ ±1‰). Changing redox conditions can partly account for ?34S variations in the tuff. Sulfur sources include magmatic rock sulfur (-1.5‰ to +9.7‰) as well as sedimentary sulfides and marine sulfate."

Jan 1, 1999

By Simon Houlding

"The objective of Part J of this paper is to identify the limitations of the traditional modelling methods and to illustrate, by example, the potential for error created by their application to complex deposits and mining situations. Part 2 describes the evolution of new, proven technology which eliminates these deficiencies and provides precise and practical alternatives. These promote the use of realistic geological and mining shapes and eliminate altogether the rectangular approximations of traditional methods.The Case for Precision ModellingIt is not the intent of this paper to promote the benefits of computer modelling per se. It is assumed that these are recognized by the mining industry in general. The concern of Part 1 is to identify, qualify and quantify (where possible) the limitations of the tractitional modelling methods. In many instances these methods continue to be applied to complex geological and mining situations for which they were never intended nor designed. The source and magnitude of the resulting errors, conflicts and inefficiencies are not always obvious, however, they can have far-reaching effects on the efficiency and profitability of mining operations.The conclusions support a rapidly growing demand by industry for practical, precise modelling technology which eliminates the deficiencies and can be applied with confidence to evaluation, design, planning and control of mine production in all situations. Part 2 describes new technology which has evolved in response to these industry pressures and has already been successfully applied to complex mining situations. This new technology can aptly be described as appropriate technology for the 1990s."

Jan 1, 1991

By M. Hozayen, E. S. Radwan, P. A. Nawrocki, A. Jebbouri

"A major geomechanical study was conducted for one of the UAE onshore oil fields with the purpose of studying reservoir rock properties and stress distribution, based on which a 3D coupled reservoir geomechanical model was built. This model was later used for wellbore stability analysis and for analyzing the effect of fractures on future reservoir production performance, including fracture deformation and permeability evolution occurring due to reservoir depletion and resulting changes in local stress field. The subject of this study was a complex fractured Upper Cretaceous carbonate reservoir characterized by high degree of heterogeneity, both vertically and laterally. Wellbore instability related drilling problems were encountered while drilling the wells in this field. The problems were mainly in drilling the horizontal wells, including breakouts and mud losses. Experimental results of laboratory tests are summarized in this paper whereas wellbore stability analysis and formulation of a 3D geomechanical model that was built for that purpose is summarized in a sister paper (Part II).A suite of tests reported includes tensile strength, uniaxial and triaxial compression and permeability under stress at simulated reservoir conditions. All tested rocks have been grouped into twelve test units and reservoir geology, rock types and challenges related to designing a viable sampling and testing program for a big reservoir are discussed first followed by sample selection criteria and details of the testing program. All test units showed brittle behavior at zero confinement and some samples showed similar behavior at low confining stresses. With confining stress increase, the behavior gradually turned to plastic but some samples showed features of ductile behavior even at low confinement.Obtained experimental data have been summarized in the form of strength criteria and correlations. The Mohr-Coulomb strength equations were obtained for each test unit and also for the averaged values. It was noticed that triaxial compressive strength (TCS) increases with decreasing porosity or increasing density but their influence on TCS is getting smaller with confining pressure increase. Good correlation of porosity and density with Young’s modulus, Poisson’s ratio, unconfined compressive strength (UCS), friction angle, cohesion and tensile strength has been noticed and there was an increasing trend between the confining pressure and TCS. Rock type (mud or grain supported) was not influencing the TCS as much as porosity and density and the two lithologies encountered affected TCS approximately in the same way. These results and correlations are summarized and explained in this paper. They were further used in the numerical modelling part (Part II) of this paper, Hozayen et al. (2015). It is also concluded that rock lithology is not a decisive factor and mechanical properties of tested rocks should be rather considered dependent on porosity and permeability. This helped greatly in reducing the number of tests that had to be performed."

Jan 1, 2015

By M. E. Sirkka

"INTRODUCTIONInco Ltd. has ten operating mines in the Sudbury district. Over the last ten years mining methods have changed at Inco with special emphasis on the introduction of bulk mining replacing conventional mechanized cut-and-f ill and undercut-and-fill methods. At the same time, the technology and size of LH D equipment has also progressed to meet the demands of bulk mining. This paper will concentrate on various aspects of the introduction and operation of electric LHD equipment a t Inco Limited, Ontario Division. BackgroundInco Limited received its first electric 2 cu. yd capacity LHD in July 1975 at Creighton Mine. The unit was tested pulling shrinkage stope muck and then was moved to a mechanized cut and- fill complex on 6600 Level. The electric scooptram received quick acceptance from the operators, due to its quiet operation and low heat generation, in a relatively hot operating environment. Although there are only two of these units operating a t present, this initial favourable experience, both in cost and efficiency, led to the development and purchase of electric scooptrams of the 5 cu. yd size. There are currently eleven 5 cu . yd electric LHDs in operation.The most recent development a t Inco is the move to yet larger electric LHDs in the 8 cu. yd size. The first electric ""Eight"" has now been in operation at Copper Cliff South Mine for over a year. This unit has the distinction of being the first one of its kind in service in North America. At the time of writing, there were ten 8 cu. yd electric ma chines operating in Inco's Ontario Division Mines, with three on order."

Jan 1, 1989

By Tony Lipiec, Ian Orford

"A key part of plant design is the selection of the design tonnage which is the hourly tonnage target that the equipment selection is based on. Its selection is critical for ensuring that the nominal tonnage which is critical to the financial performance of the project will be met. With deposits being developed at lower grades, higher tonnages and a higher degree of hardness variability, it is essential to ensure that the design process selects the comminution equipment which will be able to deal with the impact of this ore quality variation.Ore hardness and its variance are key factors to be taken into consideration for the selection of the design tonnage. It is well accepted that for the nominal or name plate tonnage to be consistently achieved, it is necessary to base the design tonnage on a hardness value greater than the average ore hardness of the deposit to be mined. The question is how much greater should this design value be to ensure that the nominal tonnage is met to the satisfaction of the project financial model. Standard industry practice is to set the design tonnage by:1. Select an ore hardness percentile that is greater than the average ore hardness to be the basis of the nominal tonnage. Typically either the 75th or 80th percentile value from the data set will be selected.2. Selection of a design factor to multiply by for determining the design tonnage, i.e. the design tonnage = tonnage (based on the average ore hardness) multiplied by design factor.This paper discusses the selection of the design ore hardness through the use of modelling or the use of a design factor. It further shows the impact of ore hardness variance on the capacity of the process and its ability to meet the nominal tonnage. Insight is provided as to what ore hardness percentile value should be selected balancing the shortfalls of tonnage for ores harder than the design value versus the above target production achieved for ores softer than the design ore hardness."

Jan 1, 2016

By S. B. Mishra

Ni-20Cr metallic coating was deposited on a Ni-based Superalloy (3Fe-19.5Cr¬0.3Ti-Balance Ni) by a shrouded plasma spray process; NiCrA1Y was used as a bond coat. Erosion studies were conducted on uncoated and plasma-spray coated superalloy specimens using an air-jet erosion test rig at a velocity of 40 m/s and impingement angles of 30° and 90°. The coating has been characterized for porosity, microhardness and microstructure and X-Ray Diffraction (XRD). An electron probe micro analyzer (EPMA) was used to obtain X-ray elemental mappings of the coating along the cross-section. Scanning electron microscope (SEM) with EDAX attachment was used to analyze the eroded surfaces.

Jan 1, 2005

By A. LA. Salama, D. B. Hyma

"The Western Research Centre (WRC) of CANMET located in Devon, Alberta, is a modern research facility offering services ranging from fundamental research to pilot plant investigations. Recently, a 0. 76 metre diameter by 7.9 metre high Flotaire column flotation cell was installed as a demonstration unit for Canadian coal producers. Because the column cell is intended to be used for research, development and demonstration, WRC recognized that additional sensing equipment and improved process control would be required to generate good data for subsequent column scale-up and engineering design.A low cost, personal computer based control system was, therefore, designed and installed. This paper describes the stepwise procedure in developing the control strategies, selection of system software and hardware, and installation and configuration of the system.Preliminary performance data on a western Canadian coal sample are also presented. Future testwork is planned on additional coal samples to expand upon existing database."

Jan 1, 1993

By P. A. Studemeister

"A trondhjemite stock intrudes a volcanic sequence that has massive sulphide lenses and that is metamorphosed to the greenschist facies at Gutcher Lake, 25 km north of Wawa, Ontario. The stock is 4 km2 in plan and is partly enveloped by an aureole of chlorite-epidote-amphibolehornfels up to 1 km wide. Within this contact aureole, biotite is partly pseudomorphosed by chlorite;epidote and hornblende are mottled by chlorite, calcite, plus quartz; albite has 0% An to 8% An; and hornblende porphyroblasts have actinolite rims. The stock has chlorite pseudomorphous after biotite; and feldspar is mottled by white mica plus calcite and has a clear rim of albite. Fractures filled with quartz, calcite, ankerite, chlorite, white mica, pyrite, and native gold cross-cut the stock and its hornfels aureole. These gold-bearing veins are enveloped by foliated rocks with quartz, calcite, white mica, and chlorite that grade outward from the veins into trondhjemite or hornfels.The stock, emplaced into a submarine volcanic series, emanated heat that caused seawater to convect and formed a contact aureole of hornblende hornfels. The stock eventually intruded its aureole and this culminated with regional metamorphism to the greenschist facies. Regional metamorphism hydrated the stock and its hornfels aureole. Native gold precipitated with quartz along fractures from metamorphic flu ids in response to cooling and wallrock hydration. exploration for this type of gold deposits should focus around felsic intrusions that have albite and that have contact aureoles retrograded by regional metamorphism."

Jan 1, 1985

By Raymond D. Potts

"Minntac is a taconite mining, concentrating and pelletizing operation located near the town of Mountain Iron, Minnesota, U.S.A. It is totally owned and operated by USX Corporation and represents its largest source of domestic iron units. Minntac began producing fluxed pellets in 1988 and has produced both acid and fluxed pellets since then. This paper compares the acid and fluxed pellets produced at Minntac. It reviews modifications that have been made to the plant flowsheet and discusses areas recognized for improvement since the beginning of fluxed pellet production. It also comments briefly on the performance of Minntac fluxed pellets in USX blast furnaces and the future of fluxed pellets at Minntac.IntroductionMinntac was built in three steps with each step designed to produce 6 million long tons of acid pellets per year. The actual production has varied though. Figure I shows the production by year at Mirmtac. Step I began production in 1967, Step 2 in 1972 and Step 3 in 1978. Mirmtac produced 16.5 million tons in 1979 but only 3.5 million tons in 1982 when the facility was shut down for six months because of the lack of demand for pellets. Since the initial start up, Mirmtac has produced over 200 million long tons of pellets. Until 1987, all the pellets produced were ""acid"" pellets. In 1987, equipment to produce ""fluxed"" pellets was installed and tested. In 1988, full scale production of fluxed pellets began."

Jan 1, 1991

By A. Mcdonald

"An X-Met Analyzer was purchased from Outokumpu in late 1983for use in the Hudson Bay Mining and Smelting smelter. Its first use was in the fuming plant to give zinc assays of each charge of slag tapped from the fuming furnaces. This quickly told the operators exactly when each charge had reached its economic tap-out point, and gave supervision a fool-proof method of identifying good operation. These and other factors led to major improvements in coal utilization in the fumers and to major increases in zinc oxide production. A second unit was later purchased for the reverberatory furnace to give silica and iron assays of the slag, and these were used to make adjustments in the feed composition to maintain the desired SiO2:Fe ratio in the slag. Since purchasing the units for the furnace, the oil consumption and copper content of the slag have dropped significantly. IntroductionSmelters have lagged far behind concentrators in process control by mean s of rapid product analysis. The on-stream X-ray analyzer was developed about seventeen years ago, and soon concentrators around the world were getting assays of the incoming ore, .concentrates and tailings every fifteen minutes, and using these to control reagent additions and other process variables. Almost always, the concentrate grade and recovery improved.In smelters, product streams are solids or molten materials, not the more easily handled slurries, and until seven years ago no inexpensive equipment was available to provide rapid analysis of this type of material. In approximately 1980, several companies developed instruments based on irradiating solid samples with low-intensity gamma rays from radio-isotope sources, and measuring the secondary X-rays emitted from different elements in the sample to provide an analysis of the elements desired.The smelter staff had a close look at what was needed and listed the following points as being critical:1. ability to provide Zn assays of a sample of slag within 15 minutes of taking it, in order to achieve optimum fuming times (zinc fuming , being a batch process, had the greatest potential benefit from rapid slag analysis);2. ability to give Cu and Fe assays on the slag, and on other smelter products, such as matte and converter slag;3. ability to give Si02 assays since the SiO2 content of the slag has a strong effect on fuming rate; and4. ability to stand up to an industrial atmosphere such as exists in the fuming plant."

Jan 1, 1987

By T. D. Traubert, J. T. Dietz, D. Catena

"Canadian process industry experts have often observed the strong trend for U.S. industrial safety and environmental laws to creep across the border. These experts will be interested in new United States (OSHA) regulations (29 CFR 1910.119), and anticipated new U.S. Environmental Protection Agency (EPA) regulations that will require over 28 000 U.S. refineries and chemical plants employing more than three million workers, to begin a comprehensive Process Hazard Analysis (PHA) of these complex plants immediately. Impending EPA regulations are expected to expand these efforts and add process Hazard Frequency and process Hazard Consequence reviews for many additional industrial locations.These regulations were created to reduce catastrophic industrial accidents, and emergency releases of hazardous chemicals which endanger employees, the public, and the environment.Accurate, and field verified Piping and Instrument Diagrams (P&IDs) are essential to permit a meaningful Process Hazard Analysis or emission release to be done. Today accurate P&IDs often do not exist. Many companies that have started to update their old P&IDs have found the process to be far more difficult, time-consuming,and expensive than they ever thought.This paper describes a new technique of computer drawn ""Intelligent"" P&IDs that can make a Process Hazard Analysis less expensive, more accurate, easier to perform, and provides lasting benefits to the plant owner long after the PHA has been completed."

Jan 1, 1993

By P. R. A. Andrews

"Talc ore from Bakertalc, High water, Quebec analyzing about 60% talc, previously dry ground to about 95% minus 53 µm was beneficiated in an 11.4 kg/ h pilot-plant flotation circuit. A concentrate of 95.9% talc (necessary for paper-grade) was obtained at a recovery at 75.3% by employing 53 g/t of Dowfroth 250 collector at pH 9.2; further cleaning increased the grade to 97.2% talc at a recovery of 57.5%.Flotation of fresh ground 60% minus 44 µm ore at a feed rate of 18.6 kg/ h produced a grade of 96.4% talc with a recovery of 78.2%, using Aerofroth 70 collector a six carbon-branch chain alcohol, at pH 9.2.The similarity in the concentrate grade and recovery between pre-ground and fresh-ground ore indicated that it was not neces¬sary to grind to 95% minus 53/2m to achieve comparable selectivity. A grind of 60% minus 44 pm corresponds to 100% minus 300 µm.Flotation retention times obtained with fresh ground ore indicate 33 min. for combined rougher/scavenger, and 15 min. each for the first and second cleaners.Talc-magnesite flotation separation is explained in terms of physico-chemical forces, with reference to surface charge and adsorption mechanisms."

Jan 1, 1989