Mining in the Arctic - History, page-2

7-93_mcneil_Nanisivik Mine.pdf

Nanisivik Mine - Operations and Innovations in an ArcticEnvironment (Old Technical Paper extract, operated by Cominco Ltd. from 1976 to2002, Mine now closed!!!!)

By W H McNeil, K R Rawling and R A Sutherland

Abstract:

The Nanisivik Mine is located 750 km north of the Arctic Circle at the north end of Baffm Island in the Canadian High Arctic, and has been in production since 1976. The paper describes the regional and local geology and the current mining and mineral processing operations, including recent innovations in each area. The Nanisivik zinc-lead deposits comprise Mississippi Valley type sulphides in a bloclc-faulted sediment sequence of Proterozoic age. The deposits formed in two mineralizing stages by replacement of the carbonate host rock. Exploration techniques adapted to the permafrost environment include high frequency EM and ground probing radar. Core drilling is done with calcium chloride brine but trials with heated flushing water indicate potential for cost reduction. The predominant mining method at Nanisivik is room and pillar. Cut and fill is used in some satellites to the main zone. Due to the permafrost, dry drilling is employed in the underground mining. The permafrost enhances ground stability, permitting large underground openings and, as a result, the use of relatively large scale mining equipment. For power conservation, a remote control system for the ventilation fans has reduced mine power costs. The milling of the simple coarse-grained ore in an arctic environment requires special techniques in the design and operation of the plant for materials handling, utilisation of waste heat, reagent selection and tailings disposal. Regulated heavy metal contents in effluents discharged to the sensitive environment are achieved by natural processes. A recent change from sub-aqueous disposal of tailings to beach type deposition within a cell in the tailings area is reviewed. Planning closure of the property presents special problems and an opportunity for innovative techniques. Approximately 25 per cent of the employees at Nanisivik are drawn from the Inuit population of the Canadian Arctic. As the first permanent industrial activity in the High Arctic some innovative approaches to employee relations were required which have resulted in a very stable, culturally mixed community in a remote location.

Introduction:

The Nanisivik Mine is located on Strathcona Sound at the north end of Baffm Island, 750 km north of the Arctic Circle, and has access to the Atlantic Ocean via Admiralty Inlet and Lancaster Sound. The occurrence of mineralisation on the shores of Strathcona Sound was first documented in 1910. A prospector, Arthur English, was taken to the site by the Inuit people of the area. English was part of the ship's company of the vessel Arctic, commanded by Captain Bernier, which was exploring the area for the Canadian government. Thepresence of pyrite crystals was well-known to the Inuit people and the area wascalled Nanisivik in Inuktitut, meaning 'the place where people find things'. No further work in the area was conducted until 1937. Two prospectors, J F TIbbiu and F McInnes, travelled by dog sled and foot to Arctic Bay from Churchill, Manitoba, a trip of some 2300 km (The Northern Miner, 1937). They arrived with only sufficient time to stake four claims in the area before leaving on the Hudson Bay Company supply ship putting into Arctic Bay. The next phase of exploration was by R G Blackadar who mapped the area around Strathcona Sound for the Geological Survey of Canada in 1954, noting occurrences of galena and sphalerite in pyrite zones.

In the late-I950s, Texasgulf Inc became interested in the mineralisation on Strathcona Sound within a major program under way in Canada to search for base metal deposits. In 1957, Texasgulf geologists, R D Mollison and W Holyk prospected in the Nanisivik area. In following seasons, surface diamond drilling and underground exploration outlined a potential lead-zinc orebody. However, the Texasgulf quest for base metals also resulted in the discovery of the major Kidd Creek deposit in Timmins in 1964. Over the next few years, Texasgulf's efforts were concentrated on Kidd Creek. Mineral Resources International Limited (MRl), a Calgary-based oil and gas company interested in investing in base metals, acquired an option on Nanisivik in 1972. A feasibility study had indicated that Nanisivik would be a viable mine, despite its location, but the industry in North America was unconvinced. With limited experience in the mineral resource business, MRl contracted with Strathcona Mineral Services Limited (Strathcona) of Toronto to undertake the management of the project. Partner funding for the mine was obtained in Europe from Metallgesellschaft AG, Germany, and Billiton BV, The Netherlands, both of whom wanted a long-term supply of concentrate for smelting. Furthermore, the Government of Canada was interested in establishing an industrial base in the High Arctic. It undertook to provide certain infrastructure and took an equity interest in Nanisivik Mines Ltd. In the 1980s, Conwest Exploration Company Limited acquired an interest in MRI and increased its interest over time, acquiring 100 per cent control of Nanisivik Mines Ltd in 1990. Strathcona has continued to provide project management services for Nanisivik. Construction started in 1974 and the official start-up date was 1 October 1976. The initial reserves were 6.3 million tonnes grading 14.1 per cent zinc and 1.4 per cent lead. At this time, over ten million tonnes grading ten per cent zinc and one per cent lead have been mined, as a result of reduction of the original cut-off grade, addition to the reserves by defInition drilling, and development of satellite ore-bodies. Nanisivik has been in continuous production since 1976 and mining operations are expected to last more than 20 years.

The Nanisivik Mine consists of a 2000 t/d underground mine and mill. Operations are continuous throughout the year. The predominant mining method is room and pillar but cut-and-ml mining is used in some satellite zones. Conventional flotation is employed to produce lead and zinc concentrates. The concentrates are stored on site over the winter until the shipping season.

The infrastructure includes the following facilities:

• community for 350 people,

• ten megawatt diesel power plant,

• recreation centre with swimming pool and gymnasium,

• day-care centre for children over three months,

• school to Grade 8, and library, nursing station with resident nurse,

• the first deep-sea port in the Canadian Arctic,

• jet airport, and

• RCMP station with two officers.

A site plan of Nanisivik is shown in Figure 2.

The weather is of considerable interest, both to Nanisivik residents and visitors. Temperatures in the winter are normally minus 20 to minus 40°C; the maximum summer temperature is plus 15°C. Due to the location 750 km above the Arctic Circle, the sun does not rise above the horizon for three months in the winter, and does not set for three months in the summer. The winds determine how long people can be outside. Storms with winds up to 100 lcmIh have been recorded. The area is a semi-desert with total annual precipitation of about 125 mm.

THE NANISIVIK ADVANTAGES

Nanisivik is a low cost zinc producer. The project has some natural and organisational advantages, which contribute to this situation.

Natural advantages

While it was the location and the weather that caused potential North American investors concern, in many ways these factors are natural advantages for the Nanisivik project:

• proximity to the ocean reduces transportation costs for concentrate shipments and incoming bulk supplies;

• the cold climate caused deep permafrost, which contributes to stable ground conditions in the mine;

• the size and shape of the ore zones, mainly sub-horizontal and up to 30 m thick, and the stable ground permit large excavations as well as use of large scale equipment underground;

• the main ore zone outcrops at both the east and west ends of the deposit; and

• the amount of waste development necessary to access the ore was minimal.

The result is a direct mining cost of approximately $ClO per tonne of ore and waste mined. Similarly, the simple metallurgy of the zinc and lead mineralisation results in direct milling costs of $C10 per tonne of ore.

Obviously, the location and climate have some adverse effects. Supplies are not merely a telephone call away. With the bulk of the required supplies arriving on the annual sea lift, the ordering of material has to be planned up to 18 months in advance. The sea lift inventories of fuel, explosives, reagents and other bulk supplies have to be financed. Perishable and high cost items are delivered twice weekly by air.

GEOLOGY

The Nanisivik zinc-lead sulphide deposits are carbonate hosted, similar to Mississippi Valley type deposits, in a block-faulted sediment sequence of Proterozoic age.

Local and surface geology

The moderately rough topography of the Nanisivik area reflects the fault scarps and is accentuated by land rebound after major glaciation. The area has poor outcrop but felsenmeer gives good definition of the geologic units. The main geologic units are the dolostone of the Upper Society Cliffs formation, which hosts the mineralisation, and the shale and dolostone of the overlying Victor Bay formation (Figure 3).

Extensive block faulting has resulted in a horst-and-graben structural setting. Mineralisation. The mineralisation formed by replacement of the host rock in two main sulphide events. An early barren massive pyrite-dolomite event, although uneconomic, was characterised geochemically by high lead-to-zinc and silver-to-zinc ratios. The later event, pyrite-sphalerite-minor galena-dolomite, produced the economic mineralisation. Nanisivik is a major sulphide deposit in terms of tonnage. Estimates range from 50 to 100 million tonnes of massive pyrite, with potentially only 15 million tonnes being economic due to zinc enrichment. The largest pyrite bodies are the South Boundary Zone and the North Pyrite Zone (Figure 4).

The deposits occur mainly in a three by five kilometre area but identical mineralisation is found 15 km and 30 km east. The deposit geometries are wedge-to-carrot shaped zones along faults, sheets along dolostone-shale contacts, lenticular zones with flat upper contacts at specific elevations, and very irregular sub-vertical zones. The Main Lens is cut by a diabase dyke of Proterozoic age and is thus pre-Palaeozoic in age, which is old for a Mississippi Valley type deposit.

Geologic model

The geologic model for the formation of the deposit invokes dewatering of basinal sediments with lateral flow along permeable clastic and volcanic units. The fluids, containing metals scavenged from the clastic rocks, rose along faults and deposited sulphides upon mixing with a reducing agent, possibly natural gas, or hydrogen sulphide. Some sort of fluid interface is indicated by the flat upper contact of the Main Lens (100 m by 3000 m) and by the restricted vertical range of sulphides at each locale. Sulphides cut horizontally across the inclined bedding within the carbonates. The main mechanism of emplacement was replacement of carbonate, not open space filling. However, some early karsting probably provided permeable openings for the fluids to penetrate. The main controls on the mineralisation are the interface of ore with reductant fluids, permeability channels within clastic units, faults, carbonate-shale contacts, and the presence of carbonate host rock in the right place at the right time so replacement could take place.

Exploration

In the Nanisivik area the rock is permanently frozen to depths of 500 m or more. The permafrost has a major impact on exploration activities. The permafrost decreases bulk electrical conductivities, but enough conductivity contrast remains between massive pyrite and wall rock to permit the use of EM geophysical prospecting methods. During early exploration,systematic vertical loop (REM) surveys discovered most of the known sulphidebodies. Turam and GEM-8 surveys have expanded the area covered and depth of penetration. VLF is the most productive method in areas of dolostone or shallow cover. Low conductivity requires high frequencies, 3000 to 5000 Hz for horizontal loop methods, which limits the usefulness of pulse EM systems such as Sirotem. One unusual technique has involved the use of ground probing radar, which was developed for use in non-conductive salt domes. This technique is usable in dolostones, but the more conductive shale absorbs the signal. A second technique has utilised portable seismic equipment. but to-date the results have been inconclusive. Geophysical prospecting has, however. beendemonstrated to detect massive sulphide bodies under the conditions atNanisivik at depths exceeding 100 metres. For diamond drilling. calcium chloride is added to the flushing water to prevent freezing. For in situ rock temperature of minus 12°C a solution of 17 weight per cent CaCl2 is theoretically sufficient However ambient temperatures in practice range down to minus 40°C and to prevent slush from forming in the tanks and piping, the working mixture ranges from 25 to 35 weight per cent CaCl2. The drill fluids are captured and re-circulated as much as possible because of the extra cost incurred in purchasing, transporting and handling the calcium salt. On more remote exploration sites. the cost of transporting the salt becomes prohibitive which has encouraged experimentation with heated drill water. A simple system of one or two in-line coil heaters burning diesel fuel to heat the flushing water appears to offer significant potential for savings. The water must be kept circulating to prevent freezing the drill string in the hole. An emergency supply of calcium chloride nearby is recommended to keep the hole open in the event of drill breakdown.

MINERAL PROCESSING AND ENVIRONMENTAL PROTECTION

Process description

The coarse-grained ore consists of massive pyrite, with a minor marcasite content, sphalerite as the principal economic mineral, and a minor quantity of dolomite, calcite and galena, with approximately two per cent ice, and an insignificant amount of quartz. The silver content of the ore is mostly included in the sphalerite. Dolomite content varies considerably with location in the current operation. A simplified diagram of the process flowsheet includes all essential features of the concentrator operations.

Crushing

The run-of-mine ore is crushed underground through a 1040 x 1200 mm jaw crusher to 115 mm size and reduced to concentrator feed size of 18 mm through a 1676 mm shorthead crusher in closed circuit with a single deck vibrating screen. After storage in an underground 5000 tonne surge bin, the fme ore is conveyed 300 m to the concentrator along a service adit.

Grinding

The 300 metre conveyor delivers the ore directly to a 2900 x 3600 mm open circuit rod mill; the slurry discharges to join the ball mill product in three 2.8 m3 unit cells, when required for flotation of a coarse primary lead concentrate. The combined mill discharge or the unit cell tailing has till recently been classified in two D15B hydro-cyclones that close the circuit to the 3200 x 3600 mm ball mill. The cyclone overflow at density 1750 g/L has a particle size distribution of 60 per cent minus 75 micrometre. The Operating Work Index of the ore is in the range of 11 - 12 kWh/tonne, but the ore is relatively abrasive. This circuit, originally designed for 500 000 tormes per year, has only undergone minor changes to achieve the current average production rate of 700 000 tonnes per year by coarsening the grind and by increasing plant availability. The system is in the process of modification to increase the production rate to 750 000 tonnes, by improving the classification efficiency and tailings disposal capability.

Lead flotation and dewatering

Lead concentrate is floated by the addition of lime, xanthate and MIBC in six 2.8 m3 rougher cells, and is cleaned in three stages of two 0.7 m3 cleaner cells, together with the unit cell concentrate if appropriate. The lime addition at pH 12.0 - 12.3 is the only depressant for the other sulphides. An experimental organic depressant DS20 was used temporarily for partially oxidised ore but was discontinued due to excessive frothing. The final lead concentrate is settled in a 7.3 m diameter thickener using a small flocculant addition and filtered intermittently on two 1.8 m vacuum discs. The fIlter cake is dried, using waste heat from the power plant, in a 1.4 x 6.1 m rotary dryer which discharges to a surge pile from which it is trucked five kilometres to the storage shed at the loading dock.

Zinc flotation and dewatering

The lead circuit tailing is conditioned in two stages with copper sulphate, incremental lime to maintain pH, and additional xanthate. Some 85 per cent of the primary zinc concentrate is floated in a bank of ten 2.8 m3 cells. The tailing is floated further with first cleaner tailing in a second rougher-scavenger bank of 14 cells to produce a final tailing. The concentrate from the two rougher sections advances to the cleaner circuit while the relatively small scavenger concentrate recycles to the second rougher. The three-stage cleaning section consists of six, four and three 2.8 m3 cells in succession which are piped for load redistribution as required. The dewatering system is similar to that in the lead circuit, using a 15.2 m diameter thickener and eight 2.7 m diameter vacuum discs.

Tailing disposal and water reclaim

The final flotation tailing is diluted with miscellaneous effluents and power plant cooling water to approximately 33 per cent solids, and pumped four kilometers through a nominal 200 mm OD pipe line to the disposal area against a 137 m static head using positive displacement diaphragm pumps. The line includes 1200 m of schedule 80 rubber-lined steel pipe followed by 2800 m of heavy-duty polyethylene pipe. By a combination of sub-aqueous and on-land deposition the solids are impounded and the circuit water is decanted and returned to the concentrator. During the annual thaw this process water inventory is augmented by the natural run-off of water from the relatively small watershed surrounding the containment area. After natural clarification of the pond water during the summer months, the excess water is decanted when it meets the quality regulated in the Water Licence for the operation. The only other fresh water added to the circuit is regulated to a maximum annual consumption, and is used for reagent mixing, domestic and minor process use, and for make-up to the power plant cooling systems.

Concentrate shipment

The annual production of some 100 000 tonnes of zinc and lead concentrate is normally transported in four ships to smelters in Europe and the southern USA. Strathcona Sound and its approaches are normally navigable and relatively ice-free from late-July to mid-October. The shipping season has been extended from late-May to late-November by the use of an ice-breaking carrier - the M V Arctic. Other vessels with reinforced bows transport the concentrate through light ice. A Canadian Coast Guard ice breaker is occasionally available to assist the carriers.

Metallurgy

Production statistics for 1992 are summarized in Table 1.

Zinc recovery and concentrate grades have not changed significantly since the start of operations. The reagent additions (Table 2) are quite simple and have minimal environmental impact.


Historically, when the ore contained one to two per cent lead, concentrates in the range of 60 to 70 per cent lead were readily produced. However the low lead content of the ore in recent years (0.2 to 0.5 per cent lead) has often resulted in lead concentrate which, not being economic, has been discarded. The simplicity of the reagent balance is also a factor in the production of low grade lead concentrate, as the high lime addition to pH 12.1 to 12.3 is the only depressant for the fast-floating, massive pyrite which is characteristic of the ore. Local pockets of bituminous ore can also contribute to problems in lead metallurgy causing a high organic carbon content in the concentrate. Spontaneous combustion of the lead concentrate occurs occasionally in the concentrate loadout, while the dryer product is still warm, apparently caused by the active pyrite and possibly the organic carbon. The combustion is readily suppressed by spreading the pile to promote cooling and compressing it to limit access of oxygen. Total consumption of wear parts and grinding media for comminution (Table 2) is quite favorable. Liner consumption in the rod mill is somewhat high which is consistent with the lime product size of 90 to 92 per cent minus 850 micrometre. Consequently the rubber liner consumption in the ball mill has been exceptional. Power consumption on the property for 1992 (Table 3) was 51 kWh/tonne of ore milled.

Of this quantity the concentrator and maintenance/services facilities consumed 28 kWh/tonne and the mine 14 kWh/tonne. Heat recovery from power generation provides all building heat requirements and the generator stack gas dries the concentrate at no additional cost. Metallurgical control was based until early 1991 on hand samples taken hourly and analysed on three generations of radio-isotope fluorescence units. As the concentrator feed grade can fluctuate hourly, the results could be misleading for reagent adjustment. In 1991 a Courier 30 on-stream analyser (OSA) was commissioned for the five major streams and used in plant control; two additional streams were added later (Figure 7). The seven streams are monitored for control purposes, but not for metallurgical accounting; the exception to date is the zinc concentrate which has been found sufficiently accurate for the daily production report. Due to the current advanced state of development of this technology, the commissioning period was minimal, and the remote location, which had been seen as a factor in not adding sophisticated technology in the plant, was no barrier to a fast interchange of information and service from the supplier. The operators very quickly adapted to the technology, and now rely on the system. The fast ten-minute cycle of dependable data, and the ability to watch circuit trends, has resulted in improvement in control and metallurgy. Logging of all instrumental readings from plant equipment is expanding to the advantage of the operators, and the manually measured densities and reagent additions are logged into the computer by the operators, for record and subsequent daily reports. The traditional operators' log sheet has been eliminated. As a result of the installation of the OSA, the Mineral Processing workforce has been reduced, by attrition, by the equivalent of over four positions.

The second phase of the OSA program, automatic reagent addition, was completed during 1993. Xanthate and copper sulphate additions have been controlled initially by an algorithm based on zinc content in concentrator feed; more in-depth relationships will be developed when a history of controlled reagent data is available. Lime addition is being controlled to a fixed total rate and later a reset may be added based on circuit pH. As pH is relatively insensitive to lime addition rate in the range 12.1 - 12.3, the improvement in control is expected to reduce the lime consumption significantly. MIBC, which is not a very significant item in plant control at Nanisivik, may be controlled at a later date.

Operating factors in an arctic environment

The ore, as mined from a permafrost environment, has a year-round temperature of minus 12°C. The ambient temperature at the plant site varies from minus 55 to plus 15°C, and the mill temperature is mostly in the plus 10 to 15°C range. These temperature differences were accommodated in the original design by the following procedure: the ore is crushed and stored underground at the ambient mine temperature, and remains frozen until delivered to the rod mill where it is slurried in water heated to 15 to 20°C. The system has worked very well over the years in avoiding the problems of freezing in surface crushing facilities and storage bins. The tailing slurry leaves the mill and is delivered to the tailing area four kilometres distant with a loss of only 2°C due to the efficiency of the 75 mm insulation on the pipe. The heat tracing on the pipe has rarely been used while operating in mid-winter, but can be activated as necessary when the line is shutdown in an emergency winter condition. Usually, the line is drained to avoid freezing. During the summer, the tailing slurry has been discharged into the pond from a floating pipe section. For some eight months of the year, the tailing pond is frozen with up to two metres of ice. Historically the tailings were deposited into the lake through a hole cut through the ice. The heat and the velocity of the tailings maintained the opening with occasional break-up of ice required by the tailings attendant. With the change to on-land deposition (see below), in winter the tailings are deposited close to the ice cover so that there is a minimal build-up of ice within the frozen tailings and so that the bulk of the water can penetrate under the ice for clarification. In the summer the discharge point is further back from the pond to maximise deposition on the beaches. The supematant water in the permanent pond proceeds under the ice to a reclaim pump system which returns it for reuse at a temperature of plus 1 to 2°C through heat-traced, insulated pipe to the mill. The tailing disposal system has been relatively trouble-free since inception, but is obviously dependent on a minimum flow of water through the circuit to prevent an excessive loss of water due to freezing in the pond. A major factor in Arctic project design is the recovery of waste heat from the diesel power generators. The significant items in the Nanisivik operation are heating of buildings and process water, and the drying of concentrate.

Waste heat is recovered from the glycol cooling system on the diesel motors through a series of heat exchangers. The entire industrial complex which is essentially contained in a single building is heated by radiators' on the cooling system. The process water reclaimed from the tailing pond is heated on entering the concentrator to absorb the excess energy, which assures that the grinding and flotation circuits operate at acceptable temperatures; the water contacting the frozen ore in the rod mill feed chute is heated to approximately 20°C. Heat in the concentrator building is also augmented by the power input to the process through radiation and steam in the flotation air, the temperature being maintained between plus 10 and 20°C throughout the year. The final concentrates from the vacuum filters are dried using stack gas from the diesel motors in co-current rotary kilns, which have no auxiliary burners. This system represents a very significant saving in process cost for fuel and has worked satisfactorily since commencement of operations, although the relatively low 370°C incoming temperature may have contributed to caking and corrosion of the kiln shells. The final moisture at five to six per cent by weight is well within the shipping regulations, and, while minimising dust, is low enough to prevent more than mild freezing during winter storage. . Approximately 12 million litres of Arctic grade diesel fuel are delivered to the site annually. As the fuel must be pumped from storage five kilometres to the plant site in winter temperatures, one grade of fuel (pour point minus 50°C) is used for power generation and diesel-powered mining equipment.

Tailing disposal

The expansion of the mine life well beyond the original projection of 12 years, based on six million tonnes of ore, has resulted in a re-assessment of future tailing disposal requirements. As originally planned, sub-aqueous deposition would result in a permanent water cover of approximately one to two metres. The original containment dam could not be raised due to the danger of contaminating the adjacent fresh water supply, so it was decided in 1991 to divide the existing area with a second dyke so that the westem one-third of the original lake area can be raised by approximately ten metres. The additional capacity created by the internal dyke should provide storage to the currently estimated end of operations. The dyke has been raised in two-metre lifts by inboard construction over existing tailings. A mixture of coarse rock and till, placed in 0.3 metre increments, is saturated with water and allowed to freeze solid during winter construction to form a permanently frozen berm which has structural strength and impermeability. The inner face of the berm is sealed with tailings after each lift. Water is currently pumped over the berm to the lower pond on a continuous basis. The tailings beaches created by deposition above the water line freeze and thaw each year. Eventually a cover of up to two metres of environmentally inert material (shale and till) will be laid over the sulphides to provide a barrier to heat penetration in summer. Testing has already shown that the penetration of frost from above and below will create a solid block of permanently frozen tailings. Test cells have been established to determine the optimum cover which will prevent the thermally active zone and oxygenated water from penetrating to the sulphides, which will thus be environmentally isolated. It is anticipated that this unique method of environmental protection after closure of the operation will be a model for other mining ventures in the Arctic where the climatic conditions are appropriate. Sub-aqueous disposal will probably be retained in the lower section of the containment area where possibly one year's tailings can still be accommodated. A study of pore water analysis in the diffusion of soluble heavy metals into the water cover will be of minimal environmental impact to the effluent from the pond after the closure of the property. The inactive pond would be a necessary part of the permanent, stable conduit, which must be provided at abandonment, for the run-off water through the upper and lower sections of the area. The frozen tailings under the protective cover in the upper section will be contoured to direct run-off water to a stable channel on the south side which will overflow to the inactive pond below.

Effluent quality.

The principal gaseous emissions from the surface complex are the stack gas from the diesel generators which present no problems, and the off-gas from the two concentrate dryers which is scrubbed through water venturis to remove particulates before discharge, in accordance with accepted engineering practice. Liquid effluent analyses are monitored regularly and the data forwarded to the NWT Water Board. These include the annual decant of excess water in the tailing system, run-off from the two open pit operations at either end of the orebody and from waste piles adjacent to the main ore zone. Natural precipitation of the lead and zinc content of the process water in the tailing pond is allowed to proceed until the regulated contents are met (Table 4).

The water is then decanted through a small holding pond where the metal content is checked on a prescribed schedule. Water analyses are done with a graphite furnace attachment to an atomic absorption spectrometer. Should the regulated contents be exceeded, the decant is interrupted until the regulations are met. The effluent is not chemically treated before release in the area of the East Open Pit, run-off water from the pit and waste piles has substantial metal content. This effluent, generated only in the summer, is collected in a holding pond and treated with lime during transfer to a settling pond. Periodically the clarified water is decanted to the local water courses after the metal content has been determined. Water accumulated in the West Open Pit is transferred to the tailing pond via the mill. During rain storms, sulphidic waste piles on surface adjacent to the mining operations add to the zinc content of the general run-off through the plant area. The run-off flows to Strathcona Sound. A monitoring program is in progress so that a plan for mitigating the effect can be developed. Since the backfill required for underground mining exceeds the quantity of the sulphidic waste on surface, it is anticipated that most of the waste will have been removed by the end of the mine life. A monthly report of effluent quality, analyses of various streams in a surveillance network, and any extraordinary discharges of plant liquid is transmitted to the Water Board routinely.

Closure and abandonment of the property

The Water License under which the mine operates requires that a plan for closure and abandonment of the property should be filed with the Water Board. This plan includes all necessary procedures to render the mine and processing areas environmentally inactive. The principal points of the plan are as follows.

1. After removal of all salvageable equipment and material, all surface buildings will be razed and the site cleaned of contaminants. Most of the waste will be burned or taken into the underground mine for permanent disposal in pennafrost.

2. The tailing disposal area will be rendered inactive by the system of aqueous and solid cover already described above to prevent oxidation of the sulphides. Man-made structures that may deteriorate will be removed and natural channels for run-off water developed.

3. Sulphidic waste rock will be returned underground as much as possible or will be used to fill the small open pits where the Main Lens has outcropped. The pits will finally be covered with inert material to promote a permanently frozen condition in the fill.

4. The mine portals will be sealed to prevent inadvertent access, and to prevent access to, or removal of, the permanently frozen material abandoned underground.

Closure of the portals will preserve the disposal facility for the discarded surface material.

Monitoring of the effectiveness of these procedures, and any necessary remedial action, will continue for some years.