The Kidd Creek orebody is a massive base metal sulphide deposit with surface dimensions of 168m by 670 m. The steeply dipping orebody has been evaluated to a depth of 1,524 m. To date, mining has been divided into three stages. The first stage was an open-pit operation that supplied ore from 1966 until 1977, when a depth of 219 m was reached at the widest, or north end, of the orebody. The second stage of mining, called the No. 1 mine, was the development of an underground mine to replace the pit and recover the ore from the floor of the pit to a depth of 792 m. Development of the underground mine started in 1969. Full production of 273,000 tonnes per month was attained in 1977. The third stage of mining, called the No. 2 mine, was initiated to accommodate a production expansion and included the development of a second underground mining operation to recover the ore from the bottom of No. 1 mine at the 2600 level, to a depth as yet undetermined, but most likely the 4,600-ft level. Development of the No. 2 mine began in 1974. Full production of 91,000 tonnes per month was achieved in 1982. The Kidd Creek orebody remains open down-dip below the 4600 level, and exploration drilling is in progress from drill stations on 4600 level. Ore-grade mineralization has been intersected to the 7,000-ft horizon.
The choice of the 2800 level as the partition between the two underground mines was based on the following factors: * The 2600 level is the lowest level which can conveniently be mined through the No. 1 mine crusher on the 2800 level.
* Development and trial stoping methods for mining at depth could be evaluated in the 2600 to 2800 block utilizing the existing hoisting facilities of No. 1 mine.
* Early development of the ore in the 2600 to 2800 block would provide additional mill-feed sooner than could be achieved by developing from No. 2 shaft.
The underground mining plan is to mine from the pit floor down to the 2,600-ft level and at the same time mine from the 2800 level down to at least the 4,600-ft level.
The Kidd Creek operations are designed to follow a logical mining sequence. The grade of ore supplied to the concentrator is the result of following the mining plan. The philosophy of mining has been to mine downward in horizontal slices. In the case of No. 1 mine, the slices are 122 m thick, while in No. 2 the slices are 183 m thick. The point of this approach is to mine the pillars immediately after creating them rather than allowing them to stand, deteriorate and generate uncontrollable situations.
Two types of ore have been handled separately prior to 1988 to optimize recoverable values in concentrates. The `A’ type ore is a copper/zinc ore with some silver and little lead. The `C’ type ore is a zinc/silver ore with some lead and copper values. The grades and recoveries of concentrates from the ores are appreciably different and make it worthwhile to keep the ores separated. The proportion of the ore types changes with depth, dictating the requirement that a portion of the concentrator have the flexibility to handle either type of ore. The mining sequence also affects delivery of the `C’ ore to the concentrator. As the `C’ ore occurs on the footwall of the north end of the orebody and the mining is retreating from hangingwall to footwall, the `C’ ore is available when it fits into the mining plan. The current mining cycle allows the concentrator to mill a single ore type, which is a blend of `A’ type ore (80%) and `C’ ore (20%) because of relatively low lead values in the `C’ ore. No. 1 mine
The Kidd Creek No. 1 mine is serviced by a 7-m-diameter shaft sunk to a depth of 930 m. A surface- connected, — 17 degrees ramp was advanced ahead of mining and currently provides access from surface down to the 4600 level. Operations in No. 1 mine are completely trackless with ready access to all working levels via the ramp and shaft. While the pit was being completed, the underground mine developed and drilled off large reserves in primary stopes from the 800 to 1200 levels upwards toward the pit floor. Development closer than 60 m to the pit floor was limited because of the effect of pit blasting on underground workings.
Upon completion of the pit, development of the crown pillar was expedited and the crown pillar was blasted into the primary underground stopes. This changed a large drilled reserve to a large broken reserve. It became apparent that the tops of some pillars were unstable because of the size and structural features, so they were reduced in height to remain stable and yet afford the opportunity to backfill in an orderly sequence, retaining fill high enough on the weak footwall of the orebody to limit failure and dilution.
To maintain a large drilled reserves, some pillars were partially drilled off with small diameter blastholes. Relaxation of the pillars made much of this drilling useless by displacing the holes to the point where they were not loadable. The pillars in many cases had to be redrilled with larger holes; this gave rise to the use of in-the-hole drills and 11-cm-diameter holes. Rotary drilling with Robbins 11D units and 20-cm-diameter holes was the most effective drilling method in the primary stopes; however, the limitations of the equipment prevented the use of these units in much of the redrilling of the pillars. A concept to mine large blocks with a minimum number of pillars is the current direction of the mine plan. Although such a method in a vertical orebody limits the broken ore reserves available, it permits the development of a large drilled reserve readily available for blasting. Mining Method
Sub-level blasthole stopes are generally 18 to 24 m wide, 30 m long and 91 m high. These dimensions are modified, where practical, because of contacts or structures. Pillars between stopes are 24 m wide and sill pillars are in the order of 30 m thick. Development of main level haulageways at 121-m intervals includes a footwall drift the length of the orebody with crosscuts at appropriate locations. A system of drawpoints on 7-m centres branches from the haulage crosscuts on either side of the stoping block. The haulage crosscuts are connected by a hangingwall return airway to exhaust ventilation raises at the extremities of the level. Orepass dumps are installed along the footwall haulageway. Sub-levels at 30-m intervals are developed from the main ramp in the footwall.
Initial development of the sub- levels is limited to that required for primary stoping, with drifting in designated pillars kept to an absolute minimum. Experience has shown that development which is convenient for primary stoping is not always compatible with the best pillar mining methods. The main purpose of the sub- levels is to serve as drilling bases for longhole drilling equipment; however, in many instances portions of the sub- levels have been utilized as production horizons to supplement capacity of the main haulage levels, to effect better draw control from stopes with mixed ore types, and to take advantage of the geometry of the orebody, particularly in the recovery of pillars.
All drift development is done with diesel-powered, rubber-mounted, 2- or 3-boom jumbos and air-powered, crawler-mounted, 2-boom jumbos. The Joy VCR-150 is the standard development drill. A recent addition to the development fleet is an Eimco- Secoma 2-boom electric/hydraulic jumbo. Development rock is mucked with 2-, 3 1/2-, or 5-cu-yd capacity LHD units. Raise development is done almost entirely with raise boring machines. Robbins 61-R units are used to drill 1.2- and 1.8-m-diameter holes. The use of a 32-R or 34-R unit for boring 1.2-m-diameter slot raises is increasing because these compact units can operate within normal drift dimensions. Stoping
Primary stopes are developed by boring a slot raise from the uppermost sub-level down to an undercut drill drift on the main haulage level. The stope is undercut in the shape of a trough and a slot is opened across the width of the stope at one end. Rings of holes, which have been drilled from the sub-levels, are blasted in vertical slices int
o the trough. The undercut is drilled with 5.4-cm-diameter holes using Gardner-Denver DH123 drills mounted on 2-boom longhole drilling jumbos. Initially, the majority of the drilling from sub-levels utilized Robbins 11-D rotary drills mounted on air-powered crawlers. These units drill 20-cm-diameter blastholes on a 6×6-m pattern. Supplementary drilling is done with Gardner- Denver DH123s drilling 5.4-cm holes or Mission in-the-hole units drilling 11.4-cm holes. The Mission and Ingersoll-Rand in-the-hole units, which drill 14-cm, 15-cm and 16.5-cm holes, now provide the majority of production drilling in both the No. 1 and No. 2 mines.
Sub-level open stoping utilizing large-diameter blastholes is the preferred mining method at Kidd Creek. The method is highly productive in terms of tonnes drilled per manshift and blasting costs, and it provides superio r working conditions. Overall dilution runs from 5% to 10%.
Development of blasting practices at Kidd Creek has kept pace with changes in hole sizes and drilling patterns. Watergels and ANFO with non-electric methods of initiation are used exclusively. Small-diameter holes are charged with pneumatic ANFO loaders or Nitro-Novel loaders in the case of watergel cartridges. Methods were also developed for loading 11.4-cm upholes, 30 m in length, with ANFO and watergels. ANFO is generally initiated with a product of Canadian Safety Fuse called Boostercord, which is a 12-grain-per-m PETN cord with 300 grain bumps every 3 m. Watergels are initiated with a product called Trunkline, which is a 150-grain-per-m detonating cord. Delays, in turn, are non-electric.
Large diameter holes initially presented a problem, with large quantities of explosives on the same delay causing severe damage to surrounding workings because of high peak particle velocities. The problem was resolved by breaking the powder column in the holes into maximum charges of 160 kg per delay. Each deck of explosives is initiated by a TNT primer on a 12-grain detonating cord with a delay in the primer. Holes have been loaded with as many as eight delays in the column. This practice has reduced the peak particle velocity to the order of 4 cm per sec at a distance of 30 m.
The mining method at Kidd Creek is an exercise in material handling. Generally, the fineness and uniformity of the material to be handled governs the efficiency of the system and dictates the degree of mechanization that can be effectively applied. The production mucking units used at Kidd Creek No. 1 mine are Wagner ST-8 scooptrams powered by GMC V-8 diesel engines. The haulage fleet is sized to operate at 67% availability, or two out of three units operating. The ST-8 scoops are standard units with up-to-date braking systems, full flow hydraulics and fire suppression equipment together with other modifications to improve mechanical availability through improved component life and ease of maintenance. A recent modification has been the conversion of an ST-8 to hydrastatic drive.
The roadways are concreted with a 35 MP a design mix averaging 25 cm in thickness. These roads are paved wall to wall with 15-cm curbs cast close to the walls. The case for concreting the main haulageways is based on the following:
* the high tread pressures exerted by fully loaded 8-cu-yd scooptrams at high speed,
* the use of copious amounts of water to control and suppress dust at the drawpoints, and
* extension of the effective operating radius of trackless haulage units because the haul portion of the cycle is done at top speed.
Concrete roadways produce excellent running conditions, but they are expensive to install and require considerable maintenance. Cleanliness is the key factor in maintaining the roadways. To this end, road cleaning equipment must be operated diligently. At Kidd Creek, fast, rubber-tired dozers are used. Periodically, haulageways have to be shut down, the roadways rebuilt and the concrete allowed to cure for a reasonable period of time. This has to be accommodated in the mining plans. Tire wear is carefully monitored as it has been proven that poor operating conditions and practices and equipment abuse are quickly reflected in the condition of the tires.
Secondary breaking is a drill-and- blast operation. Trials of mobile hydraulic rock breakers have been encouraging and appear to be worthwhile in that a significant proportion of oversized chunks can be broken in this fashion in the drawpoints. The advantage of these units is that they can operate among mucking units without the disruption and damage caused by drilling and blasting. The disadvantage is that the units are bulky and have limited reach in headings with limited back heights. The drilling units are Joy AL60 drills with feeds mounted on booms attached to the back of diesel-powered farm tractors. The booms have a hinge to permit collapsing of the unit for travelling. Sand blasts are used on chunks beyond the reach of the drilling units. The charges are placed with wooden poles or are blown through a sectional thin-walled aluminum pipe. Ground Support
The ground generally is very competent with ground support mostly a local requirement. Some cable bolting is done to prevent or delay the unravelling of shear zones, preserve strategic pillars or limit dilution. Shotcreting is done in shop areas and shaft stations. The bulk of the ground support requirement is in the development headings. Initially, mechanical anchor-type rockbolts were used exclusively. Currently, the bulk of the main haulageways, drawpoints, ramps and other permanent excavations are bolted in with grouted rebar bolts. A relatively recent development has been the use of split sets in lieu of the rockbolts in areas with a limited life span. The use of steel sets, designed in-house, has increased in recent years because of ground conditions within the pillars in the upper portion of No. 1 mine.
The centre boom of the 3-boom diesel jumbos has been modified to use the unit as a roof-bolting rig as well as a development drill. A Joy AL60 drill is mounted on a slide with cylinders to enable the drill to articulate across the width of the drift. The extensible boom gives a range of 2.4 m so that two or three rows of bolts can be placed from the single set-up. This configuration of units is, in effect, a 2-boom drift jumbo and a single- boom roofbolter on the same carrier. The difference in time required to drill off a round is seldom a factor in achieving a cycle in the Kidd Creek operations. The gains realized by mechanizing the roof-bolting part of the cycle with a minimum capital expenditure have outweighed the loss of the third drill. No. 2 mine
The No. 2 mine is serviced by a 7.6-m-diameter shaft sunk to a depth of 1,554 m. The shaft is interconnected at appropriate intervals with No. 1 shaft. Below the 2800 level, stations have been developed at 61-m intervals. A — 17 degrees ramp has been driven from 2600 level to 4600 level, with accesses to sub-levels at intervals of 15 or 30 m. The ramp was advanced ahead of mining and was connected to the No. 1 mine ramp in 1986. Operations in the No. 2 mine are trackless except for main level haulageways, where 13.5-tonne trolley locomotives pulling 4-cu-m Granby cars haul ore from transfer raises at the footwall of the orebody to orepasses close to the shaft.
The feasibility study for the No. 2 mine assumed that ground pressures would be considerably higher than in No. 1 mine and that a cut-and-fill method of mining would be used. Subsequent evaluation of residual stresses on the 2800 level, the behavior of initial development headings and the stability of large openings in the weaker waste walls prompted the trial of an Avoca type of cut-and-fill mining method. Using this method, a 9-m-wide by 60-m-high by 30-m-long stope was mined and backfilled from hangingwall to footwall in 15 m lifts. The method proved viable provided the wide entries were not required to stand too long. In another section of the orebody, between the 2600 and 2800 levels, a 15-m-wide by 30-m-long by 60-m-high sublevel blasthole stope was developed with drawpoints similar to the p
ractice in No. 1 mine. This stope was mined out and later backfilled. The excavation remained stable throughout the cycle. The stope was troughed at the bottom and peaked at the top so that the stope back could be tight-filled, except for the access drift. A pattern of stoping blocks 15 m wide by 15 m long by 200m high was subsequently laid out and a sequence for mining without pillars was embarked upon. The development of multiple drawpoints was abandoned as they would create sill pillar recovery problems at a later date.
Initially, a stope is mined out in the centre of a zone up against the hangingwall and backfilled. A second stope is mined immediately behind the first stope. Simultaneously, the stopes on either side of the initial block can be developed and prepared for mining. As the pattern develops, exposure to highly stressed ground is limited as mining along the hangingwall and retreating from hangingwall to footwall relieves the bulk of the ore zone from the high horizontal stresses. The current method is to drill off the stoping blocks from two sub-levels and the main haulage level. The undercut is drilled as a trough from the main haulage level, the first sub-level is 45 m above the haulage level.
Fans of blastholes are drilled from a single crosscut driven along the far wall of the stope towards the backfill block. The top sub-level is developed by driving crosscuts down the centre lines of the stoping blocks to the hangingwall. This sub-level serves as a base for slot raise boring, longhole drilling and the delivery drift for backfill. The stope is peaked at 50 degrees so that the fill, which runs to an angle of repose of 37 degrees , will pack tightly against the sloped backs of the stopes. A sub- level 15 m above the haulage level was driven as a conservative measure to control slot excavation and for the formation of a fill rill. Improved drilling and blasting practice has eliminated the need for the extra sub-level.
The method avoids generating wall- to-wall pillars and, except for slot mining, allows development and longhole drilling to proceed well ahead of mining, provided steps are taken to preserve the blasthole collars. The peaked back layout has been stable to date. Shrinkage of the backfill has not occurred and back conditions have been excellent. Currently, stopes are being designed with a flat mining line on a trial basis.
Slot raises are drilled 1.2 m in diameter from the top sub-level to the undercut or drawpoint level utilizing a Robbins 34R raisebore machine. Three-metre by 3-m drop raises are also used in slots 15-m by 30-m in depth. The slot in the undercut area is mined out and three or four rings of the undercut are blasted and mucked out. The balance of the slot is mined out followed by vertical slices led by the undercut. Cleaning all the ore out of the trough presents some difficulties which are overcome by driving a small auxiliary drawpoint from the adjacent crosscut and utilizing radio- controlled equipment. In effect, the method is a cut-and-fill method dealing in cuts in the order of 60,000 to 80,000 tonnes and separating the development and most of the drilling from the normal cut-and-fill cycle. It has the benefit of blasthole stoping in that ground support is not part of the cycle, and miners can work safely because no one is exposed in the stope.
The amount of broken ore reserves that can be carried is limited, but the drilled reserves can be large. Adherence to a mining sequence and schedule must be fairly rigid. The only relief or protection for ore production is to operate in several areas simultaneously. The equipment used in the No. 2 mine is similar to that of the No. 1 mine, except that it has been scaled down to operate in generally smaller development headings and stopes. The major departure in equipment selection for the No. 2 mine has been the blasthole drills, the lhd units and the use of trucks for hauling backfill. A 50-tonne Kiruna electric truck has been ordered for delivery this year. This unit will haul production ore and development waste from 4700 to 4400 level up a 10% ramp.
Drilling and blasting layouts at the No. 2 mine have incorporated the use of 11.4-cm-diameter blastholes extensively. The use of 14-cm and 16.5-cm- diameter blastholes has increased in recent years. These drills are mounted on carriers which allow the drilling of 360 degrees rings. The 20-cm holes produced by the rotary drills were found to be too large to be effectively used in the smaller stopes. Gardner-Denver DH123 drills were used initially to drill the undercut trough rings, but have been replaced with 11.4-cm-diameter ITH drills. The effective drilling range of the ITH units is considerably greater than the DH123 drills. The range and hole size have reduced the development required to mine blocks of ore.
Blasting practices in No. 2 mine parallel those followed in No. 1 mine in that blasting damage is controlled by limiting the explosives detonated per delay. Charges are decked in the blastholes or alternatively, in the case of ANFO, sodium chloride is used to dilute the charge to the designed density. LHD units in No. 2 mine are 2-, 3 1/2- and 5-cu-yd units. Initially, 2- and 5-cu-yd units were used for development production mucking. Operating experience has shown that the 2-cu-yd units are acceptable, though limited, development mucking machines, while the 5-cu-yd units are good production machines but require relatively large drift dimensions to operate within government mining regulations. It was felt that the large drifts would be a cause for concern in the future and, therefore, the 3 1/2-cu-yd units were selected as a compromise to be both effective development and production units. The trucks used on the backfill haul are 11.7-tonne rear- dump units with the boxes modified to dump cleanly in standard drill drifts. J. Eric Belford is director of mining for the Kidd Creek mines. REFERENCES Blakey, P. N., Yu, T. R. and Tansay, D. O., (1976): “Kidd Creek’s Innovative Blasthole Sub- level Stoping,” Journal of Mining Engineering, June, 1976. Yu, T. R., (1980): “Ground Control at Kidd Creek Mine,” 13th Canadian Rock Mechanics Symposium. Duke, D., (1980): “Avoca Stope Mining at Texasgulf’s No. 2 Mine,” Canadian Underground Operators Conference, Sudbury. — 30 —
Be the first to comment on "MBULK MINING AT KIDD CREEK"