35

Underhand Cut-and-Fill Mining at the Lucky Friday Mine

Clyde Peppin,* Tom Fudge,* Karl Hartman,* Doug Bayer,* and Terry DeVoe*

INTRODUCTION The Lucky Friday Mine is located on the east end of the Coeur d'Alene Mining District in the northern Panhandle of Idaho. The mine site lies just east of the town of Mullan, 10 km (6 miles) west of the Montana border (Figure 35.1). The Coeur d'Alene Mining District has produced over 31,100 tonnes (1 billion ounces) of silver, 8 million tonnes (8.8 million tons) of lead and 4 million tonnes (4.4 million tons) of zinc from 1885 through 1999. The Lucky Friday Mine has produced 6,350,293 tonnes (7,371,181 tons) of ore to produce 3,570 tonnes (114,746,000 oz) of silver, 714,500 tonnes (787,600 tons) of lead, and 89,811 tonnes (99,081 tons) of zinc. 1

Production capacity of the Lucky Friday Unit in 2000 is over 1,000 tonne/d with a total work force of 210 employees. Mechanized underhand cut-and-fill and mechanized overhand cut-and-fill mining methods are used to extract silver, lead, and zinc ore from the Lucky Friday and the Gold Hunter veins. The ore-bearing minerals in both veins are tetrahedrite, galena, and sphalerite. In 1999, the Gold Hunter vein supplied 75% of the mill feed from two stopes mined using mechanized overhand cut-and-fill and two stopes mined using mechanized underhand cut-and-fill. Ore from the Lucky Friday vein was produced from two stopes using mechanized underhand mining

produced from two stopes using mechanized underhand mining methods. The average blended ore grade was 547 gm/tonne (16 oz/ton) silver, 10% lead, and 2% zinc. In 1999, 280,200 tonnes (309,000 tons) of ore was milled and produced 138 tonnes (4,441,000 oz) of silver, 25,000 tonnes (27,600 tons) of lead, and 1,720 tonnes (1,900 tons) of zinc in concentrates.

Underhand cut-and-fill mining is the method developed and used to mine the Lucky Friday vein. This method was extended to mining the Gold Hunter vein.

GEOLOGY The Lucky Friday vein crops out as a thin, discontinuous vein in the St. Regis Formation. Between the 1800 and 2000 levels (approximately 600 m below the surface), the vein crosses into the Revett Formation and extends along strike for over 457 m (1,500 ft). Mineralization has been consistently strong at the Revett-St. Regis contact to the current mining horizon 1,800 m (5,930 ft) below the surface (Figure 35.2).

In the Revett Formation, the vein averages 1.2 to 1.8 m (4 to 6 ft) wide and dips nearly vertically to the southeast. From the surface to about the 2000 level (600 m below the surface), the Lucky Friday vein strikes east-northeast. Below the 2000 level to the 5840 level (1,770 m below the surface), the vein assumes a progressively northerly trend. The main Lucky Friday vein is constrained on each end by prominent west-northwest-trending strike-slip faults locally called the North and South Control faults (Figure 35.3). On the upper levels, mineralization was often very strong along these two controlling faults, but in the lower Revett Formation, these

controlling faults, but in the lower Revett Formation, these faults contain no economic mineralization. A stratigraphic offset of 61 m (200 ft) or less is seen across the Lucky Friday vein. The vein dips at an angle of about 85&##176;, which is about 10&##176; steeper than the host rocks. In the lower levels, the vein locally reverses dip.

In plan view (Figure 35.3), the vein is S-shaped with a sinuous middle portion that coincides with a steeply plunging anticline. The vein takes this shape because of minor folds and major faults. Several persistent faults offset the vein laterally up to 6 m (20 ft). There are numerous narrow veins in the hanging wall and footwall of the main Lucky Friday vein; however, none has proven to be economic.

Lucky Friday ore mineralogy is argentiferous galena and sphalerite. The silver-bearing mineral is tetrahedrite that forms tiny discrete inclusions in the galena. Gangue minerals include quartz and siderite. The vein is easily identified, contrasting sharply with the quartzite. Silver and lead grades are higher in the quartzite of the Upper and Lower Revett Formation. Zinc values are increasing and silver and lead grades are decreasing slightly with depth. Average net smelter return (NSR) for 1999 from the Lucky Friday vein was $88.00 per ton.

MINE ACCESS AND SERVICES The Lucky Friday and the Gold Hunter deposits are serviced from the surface through the Lucky Friday's Silver shaft. The Lucky Friday No. 2 shaft serves as a secondary access and exhaust route. The Silver shaft is 5.5 m (18 ft) in diameter and concrete lined. Total depth is 1,890 m (6,200 ft) with stations at 2800, 4900, 5100, 5300, 5500, 5700, 5900, and 6100 levels. Skip loading facilities are at the 5370 and 5970 levels.

Hoisting is done with one Nordberg double-drum, double-clutch hoist 12 ft in diameter with an 8-ft face. There are two 1,118-kW (1,500-hp) dc motors driving the hoist. Muck hoisting speed is 11.6 m/s (2,300 ft/min). Personnel are hoisted and lowered at 7.6 m/s (1,500 ft/min). The hoist was designed to hoist a 9.1-tonne (10-ton) payload from 2,286 m (7,500 ft) without a cage attached. Currently, 4-m3 (140-ft3 ) bottom dump skips are used with triple-deck cages beneath the skips. Each deck of the cage can hoist 20 people, and the decks can be removed to handle long loads. All supplies and equipment must fit through a 1.7- by 1.7-m (5.7- by 5.7-ft) opening.

Compressed air is supplied by three Ingersoll Rand XLE compressors. One I-R XHE is on standby. Total capacity is 142 m3/min at 7.6 bars (5,000 ft3/min at 110 psi). Normal usage is about 60% of that volume.

about 60% of that volume.

Pumping mine water and waste water from the refrigeration plants underground is done with Wilson Snyder plunger pumps at the 2,800- and 5,300-ft levels. Normal pumping rate is 38 to 76 L/s (600 to 1,200 gal/min). That water is made up of groundwater from the Lucky Friday vein at 9.5 L/s (150 gal/min) and water from chillers (depending on the time of the year) at 28.4 to 63 L/s (450 to 1,000 gal/min). Installed pumping capacity is 114 L/s (1,800 gal/min) with all three pumps operating at once on both the 2800 and 5300 levels.

VENTILATION Rock temperatures have been measured at 40.5&##176;C (105&##176;F) at the 4900 level and 46&##176;C (115&##176;F) at the 5900 level. To cool the working areas to a maximum of 29&##176;C (85&##176;F) in the hottest part of the year, two 230 RT chillers are required to cool the Lucky Friday section of the mine. Other chillers are installed in the Gold Hunter section.

Total mine air flow is 6,650 m3/min (235,000 ft3/min) supplied by two 186-kW (250-hp) 130-mm (5-in) in diameter vane axial fans drawing from the No. 2 shaft and a ventilation raise and one 45-kW (60-hp) fan drawing from the No. 1 shaft. The 186-kW fans operate at approximately 5.4 in of water gauge. The 45-kW fan operates at about 4.5 in of water gauge.

There are also three 224-kW (300-hp), 1,500-mm (60-in), 1,770-rpm booster fans located underground. Two of the booster fans are in series on the 4050 level of the Gold Hunter vein to draw 3,700 m3/min (130,000 ft3/min) through the Gold Hunter, and the third is located on the 5660 sublevel of the Lucky Friday to draw 2,700 m3/min (95,000 ft3/min) through the Lucky Friday.

Auxiliary ventilation for each stope is supplied by two 56-kW (75-hp), 760-mm (30-in) in diameter fans in series supplying 850 m3/min (30,000 ft3/min) to each stope. Exhaust from the stopes flows to a ventilation raise system to the 5400 level then the ramp system to the bottom of the No. 2 shaft at the 5100 level.

Cooling is supplied by one 50 RT spray cooler for each stope. Chilled water at 9&##176;C (48&##176;F) from the chiller station at the 5300 level is piped to the sublevel being operated and into the spray coolers. Warm water at approximately 21&##176;C (70&##176;F) from the spray coolers is then pumped back to the chiller station. Waste heat from the chillers is carried by water that flows through the condensers on the chillers. This condenser water is piped to the main pump station on the 5300 level.

MINE PRODUCTION The Lucky Friday vein has produced over 7 million tonnes (7.7 million tons) of ore at an average grade of 533 gm/tonne (15.49 oz/ton) silver, 10.68% lead, and 1.35% zinc. Production from the Lucky Friday vein from 1994 through 1997 averaged 544 tonne/d (600 ton/d). Production from the Lucky Friday vein in 1999 averaged 330 tonne/d (360 ton/d), which represented about 25% of total production. Most of the Lucky Friday Unit mill feed was produced from the Gold Hunter deposit located approximately 1,524 m (5,000 ft) to the northwest. Reduced production from the Lucky Friday vein after 1998 was a result of higher-value ore mined from the Gold Hunter deposit. Lucky Friday reserves are calculated to the 5930 level, with known resources extending down to the 6290 level.

ROCK MECHANICS The maximum principal stress in the Coeur d'Alene Mining District is horizontal and typically strikes N 45&##176;W. In the Lucky Friday Mine at the 5100 level, 1,554 m (5,100 ft) below the surface, the magnitude of the major horizontal stress was measured to be some 69 MPa (10,000 psi) oriented almost perpendicular to the Lucky Friday vein. This high stress combined with the brittle nature of the host Revett Quartzite results in an environment that is prone to rock bursting.

To have a better understanding of the rock burst problem, Hecla installed a microseismic monitoring system in the early 1970s. The U.S. Bureau of Mines (USBM) had previously installed a seismograph to detect and study rock bursts and large seismic events. In 1989, the USBM installed a macroseismic system to allow waveform analysis of rock bursts and seismic events. A macroseismic event is generally

and seismic events. A macroseismic event is generally considered to have Richter-scale magnitude larger than 1.0.

Prior to 1986, the mining method used was overhand cut-and-fill with raise access. Several levels were being mined at once using the overhand system. In the early 1980s, there were as many as 30 stopes in production to produce some 1,000 tonne/d (1,200 ton/d). As stoping progressed upward from level to level, the sill pillars below a level became smaller and smaller. These shrinking pillars had to support the high horizontal and mining- induced stresses. As a result of the very stiff and brittle Revett Quartzite host rock, sill pillars often failed catastrophically, causing major damage to levels and stopes.

To lessen the potential for these catastrophic failures, a new mining method was selected that eliminated the formation of highly stressed sill pillars. An underhand cut-and-fill mining method using cemented sandfill resulted from a joint research project by Hecla Mining Company, the Spokane Research Center of the USBM, and the University of Idaho. This new method shifted the high stresses to below the floor of the stoping horizon similar to a longwall in a coal mine. Hence, the acronym LFUL--Lucky Friday underhand longwall--was coined to describe it.

Because the new method no longer allowed mining to take place on multiple levels, only three or four stopes could be mined at any one time along the advancing longwall front. Therefore, each stope's productivity had to be increased from 15 to 20 tons per miner-shift to over 50 tons per miner-shift. This improvement in productivity was accomplished by the introduction of ramp access, load-haul-dump (LHD) machines, and trucks.

While the LUFL mining method results in working under an engineered and stable back, stope walls must be fully reinforced with friction bolts and chain link mesh in order to contain the effects of nearby rock bursts. Despite an increase in the incidence of seismic events with increasing depth, the LFUL mining method has greatly improved safety (Figure 35.4).

SAFETY The underhand mining method and intense attention from management and the workforce improved safety performance. During the period between 1995 and 2000, there have been two periods of more than 1 year with no lost-time accidents. Figure 35.4 shows the improvement in injury rates since 1988 and the inception of the LFUL mining method.

STOPE DEVELOPMENT Access from the Silver shaft to the Lucky Friday vein is via ramps driven at 15% grade and a 3.2- by 3.3-m (10.5- by 11-ft) cross section (Figures 35.3 and 35.5). At 27-m (90-ft) vertical intervals, a sublevel is developed 100 m in the footwall of the vein. Each sublevel provides access to mine eight 3.3-m- (11-ft-) high cuts. Development is timed so that an entire sublevel is completed as the ore is mined from the sublevel above. Normal stoping progress is eight 3.3-m- (11-ft-) high cuts per year.

Eight-hundred-fifty meters (2,800 ft) of development are required each year to develop three stopes. An access ramp for each stope is driven flat from the sublevel to within 60 m (200 ft) of the vein. In this flat section is installed a spray cooler, a small storage area, and the controls for auxiliary ventilation of the stope. A muck bay and a back section 4.3 m (14 ft) high are cut in the sublevel so that 16-ton trucks can be side-loaded with 2- or 3.5-yd3 LHDs. A ramp is driven up at +20% grade to within 20 to 30 m (60 to 100 ft) of the vein. The grade of the access ramp is established so that there will be at least 3 m (10 ft) of flat ramp before the vein is intersected. A second ramp is driven down at -20% to within 35 m (120 ft) of the vein and flattened out for 5 m, or about two rounds. When all eight cuts accessed from the sublevel above are completed, the stope crew is moved to the predeveloped access and the +20% ramp is advanced until the elevation of the of the back of the advancing ramp is at the same elevation as the floor of the last cut mined from above. The ramp is then flattened out and driven across the vein and 7.5 m (25 ft) into the hanging wall to create a muck bay. The miners in the stopes will muck the rounds blasted on the previous shift into this muck bay. Later, another crew will transfer the muck into trucks for haulage to the Silver shaft ore pass (Figure 35.6).

Mining a Cut A stope crew consists of two miners per shift responsible for the complete cycle. Mining of the cut proceeds with each miner drilling and blasting one 2.4-m- (8-ft-) long round per shift each way from the access intersection. Before beginning to muck out the round blasted on the previous shift, the miners bar down any loose fill or rock above the muck pile. Then a strip of 10-gauge chain link wire mesh is hung from Dywidag rock bolts cast into the cemented sandfill on the previous cut. This mesh provides overhead protection for the miner as the round is being mucked out. Each miner mucks the round into the muck bay in the stope intersection. Once the round is mucked out, the miners then install additional 1-m- (3-ft-) long friction bolts in the cemented fill overhead as needed to tighten up the wire mesh. The walls of the stope are also supported with 1-m- (3-ft-) long friction bolts on approximately a 1- by 1-m (3- by 3-ft) pattern. When width allows and when ground conditions warrant, more aggressive support is installed, i.e., 1.2- and 1.8-m (4- and 6-ft) long friction bolts or 1.8-m (6-ft) grouted rebar bolts. All bolt installation is done with jackleg drills. When extremely poor ground conditions are encountered or when seismic activity is very high in a stoping block, shotcrete is used to support broken fill overhead or badly fractured walls. The shotcrete is applied immediately after mucking out, then bolts and wire are installed over the shotcrete.

Once the round is supported, the next 2.4-m- (8-ft-) deep round is drilled using a single-boom, diesel over hydraulic, narrow-vein jumbo. Eighteen 38- or 44-mm (1.5- or 1.75-in) holes are used to break a heading up to 3.1 m (10 ft) wide and 3.5 m (11.5 ft) high. An additional row of holes is added if the

3.5 m (11.5 ft) high. An additional row of holes is added if the face is wider than 3.5 m (11.5 ft). The top two rows of holes intersect the loose muck laid on the bottom of the cut above. The top row is collared about 1 m (3 ft) below the bottom of the loose muck and angled up so that the holes intersect the loose muck about 1.5 m (4 ft) into the hole. Below this top row, 2.4-m (8-ft) long holes are drilled to intersect the loose muck at the end of the 2.4-m (8-ft) drill hole. Below this, four more rows of holes are drilled 2.4 m (8 ft) deep.

Paper-wrapped emulsion explosive primed with nonelectric blasting caps initiated by cap and fuse on detonating cord is used for explosive. Plastic-wrapped explosives have not proven to be efficient because the drill holes are extremely rough. ANFO is not used because of the drill holes are wet, and the mixture of ANFO with cement in the fill produces ammonia. The round is timed so that the middle hole in the top row is shot first.

Experience has shown that most rock bursts occur with the blast or within the first hour after blasting. Therefore, blasting is done only at the end of each shift. No blasting is done on shift in the Lucky Friday to minimize the chance of triggering a rock burst with a blast. Shifts are spaced at least 21/2 hr apart to allow the seismic activity to settle down before the next shift comes on.

Backfill The underhand cut-and-fill mining system depends on cemented backfill of consistent quality with uniaxial unconfined compressive strengths greater than 1.5 MPa (250 psi) and less than 4.5 MPa (650 psi), with the optimum being 3.4 to 3.8 MPa (500 to 550 psi). Fill will typically achieve a compressive strength of 1.4 to 1.9 MPa (200 to 300 psi) within 3 days. Average strengths at 7, 14, and 28 days are 1.9, 2.6, and 3.8 MPa (280, 380, and 550 psi), respectively. The fill strength is controlled by the amount of cement in the batch and the water-to-cement ratio. Pozzolith additives are being experimented with at the time of this writing to improve the flow characteristics of the paste in the delivery lines, reduce the amount of cement needed to achieve optimum strength, and improve the early strength of the fill.

Fill with compressive strength less than 1.5 MPa (250 psi) will not support itself for more than 1 or 2 days at widths over 2.5 m (8 ft). Fill with strengths over 4.5 MPa (650 psi) is susceptible to damage from seismic events and stope closure, causing it to break into large blocks. Fill with compressive strengths of 3.4 to 3.8 MPa (400 psi to 550 psi) will compress during stope wall closure and during a seismic event will fracture into smaller pieces and actually compress but not break into large blocks and fall out.

Preparation for Backfilling Once mining of a cut is complete, the stope is cleaned out, and a 0.3-m (12-in) layer of broken waste rock, called prep rock, is placed on the bottom of the stope with a 1-yd3 LHD. Dywidag bolts 2 m (6 ft) long and 22 mm (7/8 in) in diameter are driven into the loose prep rock on 1.2-m (4-ft) centers. A nut and plate are attached to the top of the bolt, and tie wire is used to stabilize the bolts vertically so they won't fall over during fill placement (Figure 35.7).

When the next lift below is mined and as each round is blasted, the loose muck cushions the fill from the blast and gives a little expansion room for the blast. The loose muck falls away from the fill and the bolts, leaving a short distance of bolt exposed. The miner can then put wire mesh over the end of the bolt sticking out of the fill, attach a nut and plate, and provide overhead protection immediately.

This preparation method is used for stopes up to 5 m (16 ft) wide. Any stope or area that exceeds 5 m (16 ft) in width will have timber supports added to the prep. The 10- by 10-in timbers are installed across the stope on 2.5-m (8-ft) centers. Holes are drilled in the timbers so the Dywidag bolts can be

Holes are drilled in the timbers so the Dywidag bolts can be installed through the timber and into the loose muck every 1.2 m (4 ft) along the length of the timber. Cable is used to sling the end of the timber and secure it to the ribs. Care is taken to keep the ends of the timbers well away from the rock walls of the stope so that as the stope walls converge, the timbers will not buckle.

A wall of either muck or timber is built at the end of the section to be filled so that the filled section will be no longer than about 60 m (200 ft). Poor-quality fill is seen when sections longer than this have been excavated

When the sand fill preparation is complete and the wall is built, the stope is filled with cemented paste backfill and allowed to cure 3 days before mining of the next cut below is begun.

Paste Backfill System The backfill plant has two independent functions. The sand plant makes the backfill sand from tailings from the flotation zinc circuit. The batch plant mixes the sand with water and cement to make a paste and pumps the paste to the Silver shaft collar for distribution to the stopes (Figure 35.8). Approximately 30% of the tailings from flotation go directly to a 9-m (30-ft) in diameter, 3-m (10-ft) deep thickener. The remaining 70% of the tailings are sent to a 25-cm (10-in) cyclone. Cyclone overflow is sent to the tailings impoundment and the underflow to the thickener. Thickener underflow at 65% solids by weight is pumped to an Eimco 3- by 4.3-m (10- by 14-ft) vacuum drum filter for dewatering. Overflow from the tub of the filter flows to an 2.4- by 3-m (8- by 10-ft) vacuum filter. The filter cake from the two filters averages 12% moisture and is transferred via 0.6-m (24-in) conveyor to a sand storage building. Backfill sand is made whenever the concentrator is operating. Table 35.1 lists the particle size distribution for the backfill sand.

The batch plant portion of the plant combines the backfill sand, cement, and water in batches that are pumped underground through a 15-cm (6-in) in diameter, schedule 80, pipe in the Silver shaft and a 10-cm (4-in) in diameter pipe on the level to the stope. The proportions of sand, cement, and water in each batch are computer controlled to ensure a consistent mix that meets requirements for the backfill needed. The dewatered sand is picked up with a bucket wheel reclaimer and dumped on a conveyor feeding a sand hopper. Cement is transferred from a silo to a cement weigh hopper. The weighed portions of sand and cement are dropped into a Nikko high-intensity concrete mixer where water is added under computer control. Each batch is mixed for approximately 2 min, then dropped into the pump hopper.

The most common batch used in the Lucky Friday is a mix with 10.5- to 11-in slump. The batch proportions are 2,558 kg (5,640 lb) of moist sand at 12% to 13 % moisture, 304 kg (670 lb) Type I/II portland cement, and 360 to 378 L (95 to 100 gal) of water. This mix yields a backfill with an unconfined uniaxial compressive strength of 3.4 to 4.1 MPa (500 to 600 psi) at 28 days.

A Schwing, positive-displacement concrete pump is used to pump the paste about 27 m ( 90 ft) vertically from the backfill plant to the shaft collar. The pump speed is controlled manually, and the process control computer starts and stops the reclaimer and conveyor to match the pump speed. The paste backfill can typically be pumped at 127 tonnes (140 tons) of paste per hour.

The 15-cm (6-in) paste fill line in the shaft is vented at the top to eliminate cavitation, which causes excessive pipe wear. The air that is pulled down with the fill has not proven to be a problem in the delivery lines and may help reduce the density of the paste and thus reduce line pressures. Lines in the stopes must be very well secured because of the high surging effect caused by the entrapped air. The point of highest pressure in the line occurs at the transition from the vertical section in the shaft to the horizontal section on the level. Line pressures of 41 to 62 bars (600 to 900 psi) are typical at this point. The pressure seen here is dependent on the length of the delivery line, the slump of the paste, and the height to which the fill is transported.

The operator in the batch plant monitors the line pressure and adjusts the mix and pump speed to maintain a safe line pressure. If the pressure at the transition rises above 83 bars (1,200 psi), the programmable logic controller (PLC) will stop the backfill pump until the pressure drops below 83 bars (1,200 psi). To reduce high line pressures, additional water may be added to the mix to reduce the slump, but additional cement must be added to maintain the proper water-to-cement ratio. The pump speed can also be reduced to lower the pressures in the delivery line.

Before backfill is mixed and pumped to a stope, flushing water is sent through the line to wet the line, ensure the line is not plugged, and provide a visual check for leaks. The stope miner notifies the backfill plant when the flushing water reaches the stope, then the batch plant is started. The flushing water before a pour is kept out of the stope by installing a valve in the delivery line just outside the stope. Telephone communication with the batch plant is very critical so the batch plant operator can judge the needs for cement and monitor progress.

progress.

When the depth of the fill in the stope is 0.6 m (2 ft) higher than the tops of the Dywidag bolts, the miner informs the batch plant operator that the stope is full enough. The batch plant then begins the shut-down sequence. The remaining fill in the pump hopper, the batch being mixed, and the batch being weighed must be completed and pumped out. Once these are pumped into the system, water is pumped into the line to flush the whole system clean. The batch plant operator informs the miner when the water is entering the system. From 15 to 30 min later, the water reaches the stope. When the miner sees flushing water at the stope valve, the valve is opened, the miner notifies the batch plant operator, and the batch plant operator turns off the water.

Experience has shown that excess water in the stope, either from groundwater or flushing water, creates very poor backfill quality. Before a fill is made, any water collected in the stope is pumped out and only enough flushing water is directed into the stope to be sure the pipe line from the flush valve to the end of the stope is clear at the beginning and end of the pour.

When a cut has been filled, the mining crew begins to drill and blast out the bottom of the access ramp beginning at a point 15 m (50 ft) from the sand wall. The gradient of the reentry is set at -20%. Within this 15 m (50 ft), the bottom has been taken up sufficiently to have a face of rock 3 m (10 ft) high when the face is just under the recently poured sandfill.

This reentry time period of 3 to 5 days combined with one or two shifts repairing ground support outside the stope allows enough time for the most recent pour to gain sufficient strength to stand up while mining under the fill.

COSTS The highest single cost item for direct stoping cost is labor (Table 35.2). Miners' pay is based on a guaranteed hourly wage set by the labor-management agreement, and incentive is paid to miners on a total-cut incentive system. A value is set for the labor to mine a complete cut, and at the end of the cut, the value is divided by the miner shifts to complete the cut. Progress payments are made at the end of each 2-week pay period based on an estimate of the percentage completed.

period based on an estimate of the percentage completed. Hours spent during the filling of the stope and repair work resulting from rock bursts or poor ground conditions are not charged against the "contract."

Development costs are based on the cost of preparing the next block of ore to be mined as the current block is being mined and the cost of accessing a cut after filling a cut. Shared services are costs such as compressed air, hoisting, pumping, and refrigeration, main ventilation system, etc., that are shared among all stopes.

ACKNOWLEDGMENTS Contributors to this section on the Lucky Friday have been Tom Fudge, Lucky Friday Unit manager; Randy Anderson, senior geologist; Steve Thomas, mine planner; Karl Hartman, ventilation engineer; Doug Bayer, senior engineer; Wilson Blake, consultant; and Clyde Peppin, engineering supervisor, all of the Lucky Friday Mine staff.

REFERENCES Blake, W., and D.J. Cuvelier. Rock Support Requirements in a Rockburst Prone Environment: Hecla Mining Company's Lucky Friday Mine. Paper in Rock Support in Mining and Underground construction: Proceedings of the International Symposium on Rock Support, ed. By P.K. Kaiser and D.R. McCreath (Sudbury, ON, June 16-19, 1992). Balkema, 1992, pp. 665-674.

Bolstad, D.D., Rock Burst Control Research by the US Bureau of Mines, Proceedings of the 2nd International Symposium on Rockbursts and Seismicity in Mines, ed. by Charles Fairhurst (Minneapolis,

by Charles Fairhurst (Minneapolis, June 8-10, 1988). Balkema 1990.

Lucky Friday Mine Staff.. Section 3, Chapter 6. In Underground Mining Methods Handbook, ed. by W.A. Hustrulid. AIME, 1982, pp. 549-557

Lucky Friday Unit, Hecla Mining Company. Policies and Guidelines Handbook. 1999.

Noyes, R.R., G.R. Johnson, and S.D. Lautenschlager. Underhand Stoping at the Lucky Friday Mine in Idaho. Paper presented at the 94th Annual Meeting, Northwest Mining Association, December 2, 1988, 14 pp.

Whyatt, J.K., T.J Williams., and W. Blake. In Situ Stress at the Lucky Friday Mine (In Four Parts): 4. Characterization of Mine In Situ Stress Field. USBM RI 9582, 1995.

1. Lucky Friday Mine, Hecla Mining Company, Mullen, ID.