Search Documents

By K. M. Sarkar

"In many plant trial s(1,2) it has been proven that very small amounts (10 to 20 ppm) of uranium dissolved in mine water can be effectively recovered by the use of ion exchange resins and this uranium recovery has many advantages. In this paper an economic analysis at different levels of uranium contamination and at different market prices of uranium are described.For this study an operating mine-mill complex with a sulphuric acid leach circuit, followed by solvent extraction (SX) process, is considered, where contaminated mine water is available in excess of process requirements. lt is further assumed that the sulphuric acid eluant containing uranium would be mixed with the mill pregnant liquor stream that proceeds to the SX plant for final uranium recovery.Recovery by Ion ExchangeMine water, in this study, is defined as the water coming from the underground mine, or mine pits, and the run-off water from stockpiles, mine shops, crushing plant , washdown areas, etc.In nature, uranium mainly occurs in two stages; reduced (tetravalent) and oxidized (hexavalent) . The oxidation of sulphides present in the orebody often causes the presence of S04- and acidity in mine water, and this acidic water readily dissolves U+6 part of the depo sit. The react ion between hexavalent uranium, which is dissolved in water as U02+ 2 cation and S04 producing uranyl sulph ate as,"

Jan 1, 1984

By Gennaro Delisle

"Les Services TMG (Niobec) currently treat a niobium ore in the Chicoutimi area of Quebec. The mineral pyrochlore is present in a carbonatite deposit which assays 5% to 7% apatite.The ore is ground to minus 65 mesh and deslimed at 10 micrometres. Part of the carbonates are floated with a fatty acid emulsion and rejected to the tailings as a first step to ensure the proper conditions for the flotation of the pyrochlore concentrate. The carbonate concentrate is 65% minus 44 micrometres and is composed of dolomite, calcite, 22% apatite and a minor amount of silicates.A laboratory investigation was undertaken to determine the feasibility of extracting apatite from the carbonate concentrate. This led to the development of an economic flowsheet to produce high-grade apatite concentrate. The apatite is depressed with phosphoric acid, and the pH is adjusted at 4.8-5.0 with sulphuric acid. An additional amount of fatty acid is added to permit efficient flotation of the carbonates. 1n a second stage, the silicates left behind with the apatite concentrate can be removed by depressing the apatite with AIC3 while floating the silicates with a secondary amine.The results of the laboratory tests indicate the possibility of producing an apatite concentrate assaying 38.3% P2O5 (83.7% BPL) with a recovery of 66.2%. Pilot-plant tests undertaken at the concentrator yielded a concentrate of 34% P205 and recovery of 76%."

Jan 1, 1981

By Francis Dakubo, Courtney Young, Avimanyu Das

"Hydrometallurgy of copper ores typically involve acid leaching followed by solvent extraction and electrowinning (SX/EW). EW is the final purification stage in which a voltage is applied across two electrodes causing the copper ions to reduce and deposit on the cathode. Applied voltages are near 2 volts with current densities around 300 amps/m2. In this study, research was undertaken to decrease the energy consumed in the EW process. To do this, various cell designs were constructed and tested in which a membrane was used to separate the cathode and anode compartments with concentrated solution from SX serving as catholyte and raffinate from SX serving as anolyte. Therefore, copper in the concentrated solution would be deposited on the cathode while ferrous in the raffinate would be oxidized to ferric at the anode. Because it is assumed that ferric acid leaching is used, resulting ferric would be recycled to the leaching process. A positive side effect of the envisioned process would be the elimination of oxygen evolution at the anode and thus the prevention of acid-misting. After a cell design was selected, it was used to test a number of variables: flowrate, concentration, temperature, ion-exchange membrane, and applied voltage. Following experiments in which variable ranges were determined, statistical modeling was conducted using STATEASE software. Optimal results indicate the voltage can be decreased to 1.5-1.7 volts with current densities of 300-400 amps/m2. Significant reduction in the energy consumption was accomplished by reducing the applied voltage. Furthermore, current efficiencies and copper production rates was increased by application of the membrane.INTRODUCTIONLeaching of copper ores is achieved using mostly sulfuric acid. The copper pregnant solution is contacted with an organic solvent or resin adsorbent which picks up copper ions as in Equation 1:"

Jan 1, 2016

By Brien A. Perry, Art Wiebert

"Many organizations which have implemented Total Quality Management (TQM) are finding that the process has either stalled or is not working so well as originally envisioned. This can be particularly evident in an engineering environment where the uniqueness of engineers and their particular culture has not been adequately considered. Winning the commitment of a highly educated group ojprojemnnJJ1s to the TQM concepts, which are often initially 5W2 as too simplistic poses special challenges.Successful TQM implementations are found among technical groups where employees have significant input into creating the new work culture, feel empowered to make changes and are well trained in teamwork. A coach who is knowledgeable in implementation problems can, when working closely with the technical teams, help them circumvent obstacles and guide them successfully through the difficult early phases. The remaining key ingredient for success is supportive management which leads the change and is both informed and appreciative of the accomplishment being made.This Paper is based on first-hand experience with common problems and results encountered in implementing TQM in an engineering environment. Described in the paper are ways of avoiding these problems through awareness of the unique nature of engineering groups, implementation planning, just-in-time training and coaching people in the TQM process. The paper concludes with a focus on the opportunities and results from a successful implementation. These opportunities and results include the benefits each engineer can realize through process thinking, empowerment and teamwork."

Jan 1, 1995

By David L. Huston, Wouter Bleeker, Bruce E. Taylor, Ron Cook, Bob Stewart, Elizabeth R. Koopman

"Abstract—Detailed oxygen isotope mapping of rhyolites in and around the Kidd Creek mine indicates that the orebodies are associated with relatively low (<12‰) whole rock values that extend up to 200 m laterally from ore. The zone of low d18O values is overlain stratigraphically by a carapace of higher (>13‰) values that extends up to 1.5 km laterally from and 300 m to 500 m above ore. Immediately above the stratigraphically highest orebody this carapace is cut by a zone of lower d18O values extending from the ore zone. Zones of higher d18O values also occur in massive rhyolite bodies that occur in the footwall of the orebodies.The zone of low d18O values partly coincides with a Na2O depletion anomaly associated with hydrothermal alteration, except that the Na2O depletion anomaly has a greater lateral extent, and the low d18O anomaly extends farther into the stratigraphic hanging wall. These data suggest that d18O mapping may detect hanging wall alteration above massive sulfide bodies more effectively than conventional lithogeochemistry. Zones of high d18O values, which define zones through which large quantities of low-temperature hydrothermal fluids have passed, may also have usefulness in exploration.Oxygen isotope variations at Kidd Creek record three stages of a hydrothermal system that evolved in temperature with time, as follows: (1) an early, low-temperature event that produced rocks with high d18O values, (2) a high-temperature event that produced rocks with lower d18O values, and (3) a later, low-temperature event that also produced rocks with high d18O values. The high d18O signature of the massive rhyolite was recorded during early-stage, low-temperature (<200°C) alteration, and was probably preserved owing to decreases in permeability associated with low-temperature silicification. The low d18O signature of cherty breccia and chloritite was recorded during the thermal peak at >300°C. High d18O values in the carapace above the deposit were recorded as the hydrothermal system waned."

Jan 1, 1995

By F. Urekar, R. B. Pankhurst

"Air spaces and air gaps in an explosives column are known to allow greater distribution and cushioning of explosive energy from the rock mass. This energy distribution is enhanced if the explosives &apos; gases can be confined in the borehole for as long as possible by plugging and stemming above the air space.Power Plug is an inflatable plug which can seal a borehole at any point to separate the stemming from an air gap and explosive toe load.IntroductionThis paper will discussion-going work by C-I -L Inc. and operators of two open pit mines to study the results of using Power Plugs:I. Pit wall stability - the authors have found the Power Plug effective for wall control blasting in open pit mines.2. Selective mining of coal seams - the authors work shows the Power Plug allows mining selective coal seams by predrilling underseam waste wedges.Pit Wall StabilityWall stability in an open pit mine is essential for safety. Good wall stability can be achieved with wall control blasting techniques which cushion the effect s of the explosives charge by decoupling the charges from the borehole wall. The usual method of decoupling a charge in a large-diameter rotary borehole is to load the explosives in tubes (Fig .I) . We now believe the air deck system is effective in sea ling a borehole above the explosive charge to create an air deck within the borehole which will cushion the blast.In a linear decoupled charge, an air annulus surrounds the explosives charge and cushions the effects of detonation genera ted gases on the pit wall. In an air deck, i.e. with the hole plugged at the collar, the gases genera ted by an explosive toe load are sealed in the empty borehole column. After a period of time, these trapped gases produce radial cracking between holes along the entire length of the borehole.It is common practice to employ de coupled explosive charge columns in the borehole to limit blast damage to surrounding rock. The exact mechanism of energy transfer in such blasts, however, is not well understood. Interruption of the continuous explosive column in a borehole by means of stemming, or air decks, may also be viewed as another form of decoupled blast. Decoupling of the explosive charge from the borehole wall affects both explosive energy release and explosion pressure developed in the borehole."

Jan 1, 1988

By Rachel S. Prudhomme, Michael L. Jeremic

"The consolidation of mine backfill by free zing is a concept which has been under study in the mining engineering laboratories at Laurentian University during the past three years. The proposed underground application of freezing involves the artificial freezing of in-situ backfill by means of a refrigerated brine solution which would be circulated through a circuit off freeze pipes. The most interesting findings to date are: (I) The uniaxial compressive strength off frozen fill is considerably higher than that of cemented fill in a ratio of 20:1 (aggregate to cement), and (2) load deformation characteristics of the frozen fill suggest that a higher ice content results in more brittle failure and that samples which have initially failed can still support up 10 one-half the ultimate load without additional failure.Although at this early stage a derailed economic analysis of the procedure has not been possible, a general evaluation of the costs involved suggests that the method could be competitive with the hydraulic placement of cemented backfill. Results obtained from laboratory investigations indicate that the freezing of backfill is a structurally feasible option .IntroductionGround consolidation by freezing is based on the principle of transformation o f water contained in an aggregate-slurry into ice, which act s as a binding agent between the bulk particles, and which results in the transformation of a granular soil into a ""rock -type"" material. The abrasive nature of the particles enhance s the strength o f the ice by providing internal friction. As a result of this internal friction, the chances of plastic flow of the frozen mass when in a loaded condition are reduced(1). T he strength of a frozen aggregate is al so dependent on the grain shape of the aggregate particles which may or may not promote particle interlocking(2) and on the temperature since, below the freezing point of the water/aggregate mixture, the strength of the frozen mass increase s with decreasing temperatures.In the shaft sinking industry, artificial freezing of loose , unconsolidated soils has been in practice for several decades. Shafts for Saskatchewan &apos;s potash industry have been sun k to depths of 3000 feet by employing artificial freezing to stabilize the Blairmore Formation(3). The common procedure for freezing ground for shaft sinking is summarized as follows(4,5): I) 150 mm diameter bore holes spaced on 1.5 m centres are drilled in a circular pattern with a radius of 1.5 m outside the periphery of the proposed shaft."

Jan 1, 1985

By V. G. Minenko, A. L. Samusev

For efficient leaching of gold from sulfide ores and the oxidation of sulfides into sulfates, it is necessary to introduce a substantial amount of oxidant - atomic oxygen, hydrogen peroxide, ozone, potassium or sodium hypochlorite. In this paper, hypochlorite electrochemical treatment is proposed as an effective method of intensification of refractory ore leaching. The objective of the study is to increase the gold recovery from refractory ores with the application of chlorinated solutions that are modified by chemical or electrochemical treatment. The advantages of the use of the modified chlorinated solutions instead of cyanide for gold-sulfide leaching are considered to be the following: a high concentration of the oxidizing agent (hypochlorite ions) in solution helps to achieve a high speed of gold dissolution in the form of chloride complexes; the possibility to produce hydrochloric acid solutions that may be used for the recovering gold by electrolysis; the possibility of the processing of carbonaceous, arsenic-bearing and others kinds of refractory gold ores. The mechanism of the interaction of the modified chlorine solutions with gold-bearing sulfides (pyrite and arsenopyrite) was investigated. The efficiency of the application of chemical and electrochemical leaching method was proved. The main factor that limits the rate of gold dissolution is a high concentration of sulfate ions (SO4 2-) formed in the solution by the oxidation of sulfide minerals. The optimal parameters of chemical and electrochemical leaching method and effective ways to remove sulfate ions were determined. The method helps to increase the level of the gold recovery up to 87% from refractory ores. The intensification of leaching by chemical and electrochemical treatment opens up opportunities for the creation of modern efficient and environmentally-friendly technologies that may be applied for gold extraction from refractory gold ores.

Jan 1, 2016

By Robert B. Parsons

"This article is based on a paper presented by Jim Scott at the 1988 annual convention of the Prospectors and Developers Association of Canada. The paper suggests a number of factors that might be taken into account in negotiating royalty terms.IntroductionRoyal ties are a part of mining life in Canada. They are used by many kinds of individuals in many kind s of situations. Prospectors commonly take royalty interests back when they transfer claims to development companies. Mining companies disposing of undeveloped (or sometimes even developed) properties are often paid at least in part by royalties. Royalties are even used by institutional investors with no other connection with the resource industry.Such different royalty holders have different investment goals, different cash flow requirements, and different risk / return preferences. These differences have given rise to a number of different royalty type s. These types range from the simplest cents-per-ton royalty (where the operator pays a royalty at a fixed rate per ton milled) to net smelter return or ""NSR"" royalties (where the royalty is based on a percentage of the operator&apos;s NSR) all the way to net profit interest or ""NPI"" royalties (where the royalty is based on the operator&apos;s bottom-line net profit),The fundamental difference between the many royalty types is in the risk/ return relationship. The cents-per-ton royalty, for example, is relatively risk-free. Of the many economic factors which affect a mine, the only one which will affect the royalty holder is the amount of ore physically milled. On the other hand, the potential for significant gains under a cents-per-ton royalty is also very limited. The royalty holder will not share in any increases in the market price of the end product or any cost reductions achieved by the producer."

Jan 1, 1988

By Pierre Gaucher

It is possible to selectively mine an orebody by combining the use of the SSW probe with any bulk mining methods as well as narrow vein exploitations. In other words, thanks to the use of the SSW probe, mining operators will be able to reduce and control dilution that is generally associated to every mining operation. Already, quick and substantial savings have been obtained. Falconbridge mines have demonstrated in Sudbury and Raglan operations, that the SSW probe can log nickel, copper and iron sulphides in blast holes. They have raised the grade and reduced dilution by outlining the limits of the ore zones before selectively loading explosives in the blast holes. By lowering the probe in a blast hole, a profile of the ore content is instantly measured and recorded. Logging cored diamond drill holes at the mine have demonstrated that the ore content (nickel, copper, cobalt, PGM) as assayed is proportional to the intensity of the probe?s electromagnetic response (EM), probably because the nickel / pyrrhotite ratio is uniform within a given orebody. The probe also measures the intrinsic conductivity of the sulphides and the percentage of magnetite. All probe readings are displayed and automatically stored at a 10-cm interval within the same reading unit. A 30-meter blast hole takes less than 2 minutes to survey. The whole instrument is light (12 kg) and very simple to use. One operator can carry the SSW system and survey either down or up-holes. The probe works from ?40°C to +50°C. Plotting the measurements of the SSW probe in blast holes (2-3-4-6-10 inches in diameter) before blasting allows to determine the exact shape of the orebody within the waste rock and uneconomic low-grade ore. Thus, only the portion of the holes that contains ore is blasted. The dilution is then much reduced and, by leaving the waste and low-grade ore in place, the mine saves on transportation, blasting, crushing, backfilling, milling, etc. Thousands of dollars are thus saved. Some savings may also be achieved during the development of stopes by guiding development drifts and by replacing diamond drilling by percussion drilling. In some circumstances, when the probe shows that some of the rich ore was not drilled-off, additional blast holes will allow to recover pockets of rich ore that otherwise would have been abandoned in the walls of a stope.

May 1, 2001

By Bo R. Arvidson

"New permanent magnetic alloys have recently opened up possibilities for development of unique mineral separators. These are now making several new applications feasible. In many cases where conventional electromagnetic-induced roll separators have traditionally been used, the new permanent magnetic roll separators may reduce both over-all capital costs and operation costs.The development of the next generation of extremely powerful permanent magnets, i.e. neodymium-iron alloys, is progressing rapidly and will affect the roll separator as well as another type of magnetic separator. Permanent Magnet Roll SeparatorDesign and FeaturesDiscs of permanent magnets interleaved with soft iron discs constitute the basic magnetic roll design (Fig. I). By enveloping the magnetic roll with a thin abrasive resistant belt (Fig. 2), even highly magnetic particles are removed.The first prototype roll utilizing the new magnets was shown to have approximately equal strength as induced magnetic roll (IMR) separators. Since the initial test roll, the new separator, called the PERMROLL® , has been made twice as strong with regard to magnetic force, that is, the magnetic flux density is determined to exceed 1.7 tesla (17 kilogauss), while the magnetic field gradient is more than twice the gradient than can be practically achieved with IMR."

Jan 1, 1985

By Zia Pirzada

"Time is of the essence" and "Time is money" These are powerful and true statements during a project life cycle and have far-reaching implications on the success and failure of the project Time and progress management is crucial to the project success. Project delays may increase the cost to an unacceptable level and may render the project as economically unfeasible. On the other hand, if the project is completed on time or earlier than as planned, it may generate economic benefits greater than what had been envisioned To take advantage of the financial opportunities and to improve the cash flow as well as Return On Investment (POI), project schedule can be accelerated to complete the project earlier than planned Project schedule acceleration can be achieved by using analytical tools, such as "Schedule Optimization", "Schedule Reduction", "Schedule Compression" or a combination of any of these concepts along with the "Planning for start up". Schedule optimization does not entail any extra cost, however, other acceleration methods, such as Schedule Reduction and Schedule Compression, may require extra expenditures from the owner. Based on the forecasted early start of the revenue stream, the extra expenditure could become a worthwhile venture with a very high rate of return To accelerate the schedule, all aspects of the project should be reviewed with acceleration costs calculated. All risks and opportunities should be evaluated and taken into consideration to determine if the acceleration is feasible and advisable and will provide definite financial advantages Sometimes schedule acceleration may be desired to prevent the company from paying liquidated damages. This is especially true when the project is not expected to complete on the contractual completion date and when the cost of acceleration is less than the liquidated damages that may accrue against the company.

Oct 1, 2010

By Z. D. Hora

Commercial production of barite commenced in British Columbia in 1940; shipments have been recorded from ten deposits, of which four are particularly significant: The Brisco, Silver Giant, Parson, and Mineral King mines. Parr of the barite production is a by-product of base metal deposits in which barite is a major gangue mineral. About one hundred occurrences are known. COMMERCIAL production of barite in British Columbia dates back to 1940. Production volumes rose and fell and reached 39,900 tonnes in 1972. The main producer is Mountain Minerals Limited . There are two principal areas of barite deposits in British Columbia: ( I) The southeast part of the province along the Rocky Mountain Trench , which is the source of all production to date. (2) Stratabound occurrences in northeast British Columbia in the Gataga-Ketchika rivers and Muncho Lake areas which are not yet developed. Of the approximately one hundred reported barite occurrences, only ten have recorded commercial shipments, some of insignificant amount. So far, four deposits have proved to be important producers: Brisco, Silver Giant, Parson and Mineral King. The Mineral King property is situated near the confluence of Toby and Jumbo Creeks, 35 km southwest of lnvermere. The lead-zinc-copper-iron-barite deposits, as described by Fyles (1959, 1962), replace Proterozoic Mount Nelson dolomites of the Purcell system on the limb in the trough of a complex, tight syncline. The mine operated from 1954 to 1967 and 2,097,900 tonnes of baritic ore that was processed for base metal concentrates and also produced 22,680 tonnes of by-product barite. Since 1970, some 136,000 tonnes of barite have been recovered from the old tailings pond. The deposit is now considered depleted. The Silver Giant mine is located 10 km northwest of Spillimacheen. Barite and lead-zinc sulphides replace Cambrian Jubilee limestone near their contact with the overlying Ordovician McKay slate. Mineralization is located in the nose of an overturned plunging anticline. Between 1951 and 1957 800,000 tonnes of baritic ore were processed for lead and zinc concentrate. After 1959 barite was also produced , first from open pit, and later from reworked tailings. Up to 1979 production of barite totalled approximately 180,000 tonnes. Two other producing mines in the area are descri bed by G.R. Manson (this volume). Large, bedded stratabound barite deposits are reported from northeastern British Columbia. Six major shale-hosted Industrial Minerals in Canada lead-zinc-barite occurrences are known in the Devonian Gunsteel Formation. These are: Driftpi le Creek, Mount Alcock, Cirque, Pie, Fluke and Elf. In addition, numerous occurrences of strati form blebby or nodular barite and locally massive white crystalline barite are known from the Driftpi le Creek and Ak ie River districts in Gunsteel shale and from the Kechika, Road River and Ware map areas from Silurian siltstone . The size of these deposits is impressive; the baritic horizon at Mount Alcock is reported to be 25 to 30 m thick , and dri ll ing results from the Cirque deposit indicate reserves of 30 mi llion tonnes of baritic lead-zi nc-si lver ore. Similar bedded barite deposits are also reported from Midway area, near the Yukon border.

Jan 1, 1984

By Jorge Lema Patino

"The purpose of this paper is to determine the best techno-economic alternative for the treatment of tin ores, when comparing the various mining and metallurgical methods of preconcentration, concentration, volatilization (fuming) and up to the final stage of selling to the smelter. The study of the influence of the different parameters of cost, recovery and quoted prices has motivated the paper, which aims at determining the best process of ore treatment in terms of its techno-economic feasibility.Method of WorkOn the basis of head grades of 0.5, 0.75 and 1.0% tin, the following alternative methods of treatment are possible:1. Mining-Concentration-Smelting.2. Mining-Preconcentration-Fuming-Smelting.3. Mining-Concentration-Separation, with direct smelting and fuming/smelting of the two fractions produced.The calculations were carried out on the basis of existing information and publications concerning the various experimental recoveries in the operating mills of COMIBOL. In the case of concentrates containing less than 10% or more than 40% tin, extrapolation was required.The final grade of the tin concentrates and the gross weight of the concentrate were calculated on the basis of the estimated recovery curves. The costs of the preconcentration and the concentration stage s were determined per gross metric ton of material treated.The calculation of the cost of the smelting stage is similar for all the alternatives, the royalties and smelting charge varying with the grade of the final concentrate. In the fuming stage, a uniform recovery of 90% was taken in all cases so as to obtain a final concentrate grade of 65% tin, regardless of the head grade of the feed material.For the separation stage, the following considerations apply:a) Separation can be carried out by gravity separation or flotation methods.b) The cost of separation varies depending on the system used.c) The distribution of tin in the two fractions is 100% at this stage. In other words, there is no loss of tin during separation."

Jan 1, 1982

By Errol Ladner

The application of large-diameter bore holes in underground operations at Heath Steele was first tried in February, 1975. Since that time, 90,000 feet of 6-l/2-inch holes have been drilled, primarily in open blasthole stoping, for a total of 1.4 million tons, 37% of underground production. This paper deals with a review of drilling with conventional percussion longhole drills, a practice which is still utilized at Heath Steele, and the changeover to accommodate the use of larger blastholes in underground operations. The discussion will include layouts for 2-inch-diameter blasthole stope blocks, 6-inch-diameter stope blocks, trials for I. T.H. (in-the-hole) stope rooms, and drill pat/ems, drilling and performance. The complexity involved in the development, drilling and blasting of 2-inch blastholes is overcome by the over-all simplicity of drilling and blasting in I. T.H. stopes. This has made the practice of open blasthole stoping more economical and efficient. The future indicates an increase in the company&apos;s I. T.H. program. Investigations are being undertaken regarding longer holes to decrease development costs. A VCR stope is also being developed for trial.

Jan 1, 1979

By Y. Kocaefe, D. Bhattacharyay, D. Kocaefe

The carbon anodes used in the electrolytic production of aluminum are made of petroleum coke, recycled anodes and butts, and coal tar pitch. Due to technological advances in the aluminum production technology, aimed at increasing production and decreasing energy consumption and greenhouse gas emissions, better quality anodes are required. Presently, the anode quality is evaluated by visual inspection and the analysis of a small core taken from 1.5-2% of the anodes produced. The visual inspection is susceptible to human error and limited to surface imperfections. The analysis of cores takes time, and therefore, it is difficult to determine the existence of a problem and intervene at the right time. In addition, the core does not represent the whole anode due to the non-homogeneity of anodes. A validated online anode quality control system can highly improve the situation. In this article, an electrical resistivity measurement method developed for this purpose is presented.

Jan 1, 2014

By James J. Roettger, Robert J. Mitchell

"The results of a series of model stope pours designed to investigate working pressures on bulkheads during hydraulic pouring of classified tailings backfill are reported. Total bulkhead pressures and fill water pressures were monitored under the condition of a fully saturated backfill. Similitude of the model testing is discussed and equations are developed to estimate bulkhead pressures under prototype conditions. The model results show that good base drainage, combined with a bulkhead setback, can reduce the expected bulkhead pressures to values such that timber or concrete block structures should be capable of supporting a saturated backfill. Introduction Bulkheads are used in underground running operations to block drift openings in order that hydraulically placed sand backfill can be contained in stope pours. Whenever possible, the sand fill is designed to have sufficient percolation rate so that excess hydraulic transport water will gravity drain out through the bulk headed drifts. This avoids costly decanting systems. Early bulkhead design assumed that the bulkhead pressure could be as high as 0h = Hy where H is the fill height and ""y is the saturated unit weight of the fill (Thomas et al., 1979)(1). A typical design pressure for a 50 m fill height would be 1250 kPa giving a design force of 12,500 kN on a bulkhead located in a drift of 10m2 in area . Typical bulkheads were 1 m thick heavily reinforced concrete structures formed and poured into a notched drift. Not only were these expensive structures but the associated problems of inside drainage systems and curing time delays increased production costs. As experience with bulkheads for large bulk pours increased and fill pressure measurements became available [see, for examples, Libby et al., 1975(2), Askey et al., 1978(3)] , &apos;free drainage&apos; timber and concrete block bulkheads were designed. Substantial savings have been realized using these types of bulkheads but greater engineering design and monitoring of backfill pours is necessary to ensure that backfill pressures do not exceed the bulkhead design pressures.The effective and total stress components of fill pressure, the engineering concepts associated with the &apos;free drainage&apos; design and some detailed design considerations have been discussed by Smith and Mitchell (1982)(4) . It is important to the safe use of these designs, however, that the basic engineering concepts be demonstrated and communicated to those responsible for the day-to-day backfill operations. This paper describes and explains observations from a series of model studies designed to measure bulkhead pressures in a saturated fill."

Jan 1, 1984

By R. A. D. Pattrick, D. A. Polya, D. J. Marsh, J. -J. Bao

"The utility of high-density stream sediment and stream water sampling to the exploration for silver deposits was investigated by the study of the epithermal Ag-Au base-metal prospect at Lagalochan in western Scotland. The geology of Lagalochan is dominated by a calc-alkaline Caledonian intrusive complex emplaced into Dalradian Supergroup rocks (Harris et al., 1988a). The complex consists of subvolcanic porphyrytic rocks associated with intense hydrothermal alteration and brecciation. Early Cu-Ag-Mo-rich mineralization is found in the core of the complex with later Pb-Zn-Ag-Au-As-Sb-rich mineralization toward the margins (op. cit.).Stream sediment sample analysis spatially resolved the Cu-rich and Ag-rich mineralization. Ag and Cu anomalies relative to local background values were found in all size fractions analyzed (-125µm, +125 µm, +180 µm, +250 µm) but the more pronounced anomalies were found in the -125 µm fraction. The coarser fractions generally revealed more erratic but longer dispersion trails. The separation of Cu-rich and Ag-rich mineralization is also apparent in the distribution of relatively high dissolved Cu and Ag in stream waters, particular when expressed relative to the solute load as Cu/TDS and Ag/TDS.The mineralization at Lagalochan is indicated by Ag, Cu and Pb stream sediment and water concentrations that are anomalous with respect to local backgrounds, but are difficult to distinguish from regional backgrounds. This emphasizes the further potential of high-density stream water sampling,in particular, as a viable geochemical exploration tool on a sub-regional scale for epithermal depositsin relatively unpolluted upland terranes."

Jan 1, 1995

By S. Zador

"IntroductionIn the past decade, metallurgical and mining companies have shown an increased interest in the licensing of technology. When licensing technology, it is usually to the benefit of the parties, to have at least some of the technology to be licensed protected by a patent.This paper is intended to bridge the gap between the engineer in the field of mining, metallurgy and mineral processing who is interested in exploiting an invention commercially, and the patent agent or intellectual property lawyer whom he/she might retain to prepare a patent application and then prosecute the application to obtain a granted patent. It is also intended to encourage the inventor or a small business concern to consider the value of patents in the sale and licensing of technology. The lawyer may subsequently advise and assist the inventor in licensing his/her invention.A brief outline of some of the basic features of the patent system follows. The nature of the information, which the patent professional requires from the inventor to enable him/her to draft a patent application, will also be considered in general terms. Furthermore, various types of licenses which may be drawn up between a licensor of specific technical information and a licensee, will be listed and briefly discussed. No attempt, however, will be made to give a comprehensive and detailed account of the above aspects; the interested reader should consult legal publication and reference books, and patent professionals for further specific information.A granted patent is a document which describes an invention and how to practise it. The claims of the granted patent, figuratively speaking, fence off the intellectual property the inventor regards as his invention. The Commissioner of Patents, representing the Government, grants to the inventor or his assignee the exclusive right of making, constructing and using the invention, and selling it to others to be used, for a limited period of time, in exchange for disclosing the invention to the public. Thus, a patent prevents others from practising the claimed invention but it does not, necessarily, bestow the right to the patentee to practise his own invention. The term of a patent now in Canada and in most of the developed countries of the world, is 20 years from the date of filing the patent application. In the United States the patent term is still 17 years from the date of issuance of the patent. In Canada, the term was changed to 20 years in the new Canadian Patent Act which came into force on October 2, 1989. Any patent application filed prior to that date or granted before it, retains a 17 year life, commencing at the time of grant."

Jan 1, 1991

By K. Enns, J. J. Byerley

"While thiocyanate (SCN-) is used in a number of industrial processes (production of polyacrylonitrile filaments, synthesis of agricultural chemicals and certain metallurgical solvent extraction and flotation circuits) it is recognized more as an undesirable by-product associated with &apos; the metallurgical industry. Thiocyanate is present at low concentrations in some waste streams from the iron and steel industry. Of particular significance is the formation of thiocyanate during cyanidation of gold and silver ores, concentrates, tailing s and residues. The thiocyanate levels present in the resulting effluents may range from a few mg/I to several thousand mg/I.At present, thiocyanate does not appear to be as hazardous in the environment as cyanide (CN-) at the same concentration although its interaction with other species and resulting decomposition of products may be unacceptable. Of course the presence of thiocyanate in the leaching circuit and ultimately in the effluent streams may add considerably to the oxidative requirments of any non-selective treatment process for cyanidation wastes.With respect to the thiocyanate as it relates to the cyanidation process, two aspects are evident. Thiocyanate forms and persists in the cyanidation circuit resulting in a ""fouling"" condition with subsequent loss of leaching activity. In order to reduce or control this problem, it is common practice for some mills to discharge daily as much as 20% of the barren leach liquor. Besides contributing to fouling of the leach liquor the formation of thiocyanate represents an irreversible loss of reagent cyanide. It is noteworthy that in some present or proposed mill practice reagent cyanide loss through thiocyanate formation may reach 50%(1). The purpose of this paper is to discus s the application of controlled electrooxidation of thiocyanate with the view of suggesting a process for regenerating cyanide from waste thiocyanate(2). The paper describes the important theoretical aspects and presents experimental data obtained for thiocyanate electrooxidation using both small and large scale electrochemical reactors. Some economic conclusions are presented."

Jan 1, 1984