Rock mechanics
of 8
All materials on our website are shared by users. If you have any questions about copyright issues, please report us to resolve them. We are always happy to assist you.
Related Documents
  Jour nalPaper  Introduction Location  Konkola No. 3 Shaft is one of the two shafts atKonkola Mine operations. It is the mostnortherly shaft of the operations owned byKonkola Copper Mines Plc in the richCopperbelt Province of Zambia. The Konkolamine is located in the town of Chililabombwesome 450 kilometres northwest of Lusaka. Geological setting   The area lies on the nose of the KafueAnticline, the dominant regional structuralfeature on the Zambian copperbelt trendingsouth- east north-west. The stratigraphyranges from the Achaean Basement complexconsisting of granites and schists to the LatePrecambrian Katanga System, a sedimentaryseries containing quartzites, conglomerates,sandstones, siltstones, dolomites andlimestones.  The No. 3 shaft exploits the KirilabombweNorth orebody on the nose of theKirilabombwe anticline. The ore is hosted in ashale package comprising units of thicklybedded siltstones and laminated micaceoussiltstones with dolomite bands. The thicknessranges from 6 m on the flanks of the fold to 15m in the fold nose area. The dip varies from alow 10 degrees in the nose area to about 60degrees on the flanks. The orebody has a totalstrike length of over 6 kilometres and extendsfrom the sub-outcrop and is open ended belowthe 4 500 level. Geotechnical characteristics   The No. 3 shaft is an intermediate depth (lessthan 1 000 m) operation. In situ  stressmeasurements indicated that the maximumprincipal stress aligns with the Kirilabombweanticline axis and the Intermediate and MinorPrincipal Stresses have similar magnitudes andtrend parallel to cross faults. The k-ratio isestimated at 0.85 (RMT 2001). The rock mass condition in thestratigraphy ranges from very good in thefootwall series and the immediate hangingwall to poor and generally fair in the orebody. The rock mass condition also shows variationfrom poor in the fold axis area to fair and goodtowards the flanks. Brief mining history  Mining operations at Konkola No. 3 Shaftstarted in the late 1930s. The principal miningmethod was sub-level open stoping usingscraper units to load ore into boxes tolocomotive tramming systems. This methodwas used from the 300 Level to the 1660Level. The method was generally expensiveand of low productivity. Mechanized miningwas introduced in the mid 90s in order toincrease production in the substantial wide flatdipping resources. Post pillar cut and fill(PPCF) was used on the 1850 level advancingup dip to the 1660 level. The method wasdiscontinued because of insufficient backfill tosustain the required production rates and wasreplaced by the over-cut and bench (OCB) inlate 2000, initially without backfill but laterbackfill was introduced. Geotechnical considerations in thedesign of the MOCB mining method atKonkola No. 3 shaft by M. Lipalile*, A.W. Naismith † , and A.B. Tunono* Synopsis Konkola No 3 shaft experienced severe mining-induced stress andtime-related deterioration in ground conditions in an area where anover-cut and bench (OCB) mining method was used to extract ashallow dipping wide orebody. This caused an unacceptable risk forsafe mining operations to sustain the required levels of production.After a review of the conditions it was decided to change the miningmethod to CASCADE to recover ore in the affected area and modifythe OCB method and use a modified over-cut and bench method(MOCB) for new mining sections. This paper describes the geotechnical criteria included to derivethe mining standards for the MOCB. The geotechnical risks areidentified and taken into account to develop stope designparameters, stoping sequences and support requirements. Amonitoring programme to assess conditions as mining progresses isalso discussed. * KCM Zambia.†SRK, Johannesburg ©The South African Institute of Mining and Metallurgy, 2005. SA ISSN 0038–223X/3.00 + 0.00. This paper was first published in the SAIMM Conference, Base Metals, 26–29 June 2005. 607 The Journal of The South African Institute of Mining and Metallurgy  VOLUME 105 NON-REFEREED PAPER OCTOBER 2005  s  Geotechnical considerations in the design of the MOCB mining method at Konkola No. 3 Mining-induced stress-related damage During mining of OCB stopes between 1660 and 1850 levelsground conditions deteriorated with increased severity of damage to excavations ranging from minor blocky falls at junctions in the early stages to severe pillar sidewall spalling,hangingwall collapse and floor heave in the later stages. Thedeterioration was attributed to (Naismith et al  . 2004): increased stress levels in remnant areas as miningprogressed up dip, leaving a progressively smallremnant pillar towards 1660 level   delayed backfill placement allowed pillars to deteriorateand soften, thereby shedding load to adjacent stiff remnant areas   rapid oxidation and weathering of the orebody led toreduction of the rock mass strength.   laminations and weak partings in the orebody createdthin beams that break when left in the roof of theexcavations. Based on hazards identified, which include:   out of sequence mining   delays in installing secondary support   delays in placing backfill   weak ground that deteriorates rapidly with time andexposure to atmosphere   large excavation spans.It was decided that future mining methods and strategyshould incorporate the following:   reduce remnant areas by mining to strict sequence   have rapid support installation and backfill placementpotential   rapid extraction rate to reduce exposure time   avoid leaving thin beams in the immediate hangingwall   site major access ways in competent footwall seriesrocks   keep development dimensions as small as possible. Modified over-cut and bench (MOCB) method  The modified over-cut and bench method (MOCB) appears tosatisfy the above requirements and is being tried between1850 L and 2200 L in a block 400 m long along strikebetween 2160 m W Section and 4275 m N Section. The trialis to be extended to a second block that is 500 m long alongstrike from 4050 m N Section to 3500 m N Section. Mining layout and sequence  The major difference between the conventional OCB methodand the MOCB is that the over-cuts are aligned parallel to anapparent dip of 7°instead of being flat and along strike. Theconceptual layout is shown in Figure 1.Initial access into the orebody is via an access cross-cutfrom a waste ramp. Once in the orebody an incline (oreaccess ramp) is developed against the hangingwall onapparent dip of 7°. From this access ramp, extraction drivesare developed parallel to the strike at 100 m intervals. Thesedrives define the top and bottom of the mining block. A blockconsists of a series of 4 m wide x 4 m high over-cut raises at14 m centres mined parallel to the Access ramp leaving a 12m wide longitudinal pillar between adjacent raises. A 10 mpillar is left between the access ramp and the first over-cutraise in the block to protect the ramp and also between theExtraction drive and the bottom of the stope panels. The finalpanel spans are 10 m, separated by 4 m longitudinal pillarswith dip lengths of 80 m to 100 m. Rockmass condition in the trial area  The sequence of stoping consists of developing the over-cuts. These are slyped to a span of 10 m. Once two adjacent over-cuts have been slyped, the first over-cut raise is benched tothe footwall contact. Backfill is then placed immediately afterthe benching phase has been completed. After filling the firstover-cut, the adjacent over-cut is benched and slyping in  s 608 OCTOBER 2005 VOLUME 105 NON-REFEREED PAPER The Journal of The South African Institute of Mining and Metallurgy  Figure 1—MOCB layout  next over-cut commences. The estimated stand-up time forthe ledged over-cut before filling is 2 to 3 months. The trialarea lies in the axial fissure zone and the rockmass isaffected by joints associated with a major fold axis and someminor faults. Two prominent joint sets and two sets of random joints exist (Table I). The critical rock units are thehangingwall quartzite (roof spans) and the ore shale (benchsidewalls). From borehole information, the condition of theHangingwall quartzite and ore shale vary, as indicated in Table II. Geotechnical risks in the MOCB design  The geotechnical risks to be considered are as follows:   stability of the roof of fully ledged over-cut (10 mwide) and full height sidewall against total collapse   stability of bench sidewalls during benching andlashing operations   localized wedge and slab failures   regional stability with particular regard to out of sequence mining and security of main accesses   pillar performance during the various stages of thestope life   backfill placement—quality , scheduling and bulkheaddesign. Stability of spans Analysis   The stability of the roof span of the full width over-cut wasanalysed based on the hydraulic radius or shape factor andMathews/ Potvin’s stability number (Potvin et al  . 1989). Q-rockmass rating parameters for the hangingwall quarzitewere used because the roof of the over-cuts is designed to bealong the geological hangingwall (GHW). The hydraulic radius values are used to estimateunsupported apparent dip span limits for 10 m and 12 mwide stopes for a given rockmass condition. The results areshown in Table IV. Conclusions   The proposed 10 m wide over-cuts are stable without supportover 40 m apparent dip spans in good ground and 15 mspans in fair ground. This is confirmed by observations madein existing sub-level open stoping where stopes of similardimensions were used successfully. Experience from existingOCBs indicates that the sidewalls are stable over the fullthickness of the orebody and do not influence span stability. Localized wedge and slab failures Analysis  Occurrence of localized wedge and slab failures in the roof and sidewalls as a result of intersecting discontinuity planeswas investigated using ‘UNWEDGE’, an equilibrium stabilityanalysis computer package. The results are indicated in  Table V. Conclusions   The analysis indicates that kinematically feasible wedgesoccur both in the roof and sidewalls of the excavation. Twotypes of wedges are formed in the roof. The first type is Geotechnical considerations in the design of the MOCB mining method at Konkola No. 3 Jour nalPaper  609 The Journal of The South African Institute of Mining and Metallurgy  VOLUME 105 NON-REFEREED PAPER OCTOBER 2005  s Table I  Discontinuity sets SetDipDip directionSpacingApertureInfillSurface condition Bedding plane15–20°260°0.15–0.20 m<1 mm–5 mmKaolin FeOxSmooth-planarJ160–80°010–025°0.5–1.25 m1 mm–2 mmSilt/cleanRough planar– rough undulatingJ1c75°170–195°J270–90°070–080°0.5–1m<1mm–3mmClean, FeOx and SiltRough planarJ2c80°225–250°J375°135°1.5–2m3mm–5mmsiltRough planar Table II  Rockmass rating Rock typeThickness (m)RatingNotesQ-ratingRMR-ratingDescription Hangingwall > 204. 6Fair Massive quartzite, some kaolinization in bandsquartzite30Good to very goodMassive quartzite, minor kaolinization in bands1.1PoorQuatzite with kaolinized bedding planesOre ShaleUnit E1.560Fair to goodMicaceous siltstones interbeded with sandstone layers, kaolinized in placesUnit C/D3–4.540FairThinly bedded siltstones with dolomitic interbands, kaolinizedto give weak parting planesUnit B375Good to very goodThickly bedded to massive siltstoneUnit A0.5–110Very poor to poorPoorly consolidated sandy siltstone  Geotechnical considerations in the design of the MOCB mining method at Konkola No. 3 formed by the intersection of the bedding plane with any twoof the joint planes. These generally have a short apex length(<3 m). The second type is formed by the intersection of thetwo major joint sets and the random joint set without thebedding plane. These wedges are very large and generallyhave large apex lengths (>9 m). The failure mechanism iseither falling under gravity or sliding along one or both of the joint planes. Wedges formed in the sidewall either topple orslide along the bedding plane. In addition, observation inexisting over-cuts indicates that severe sloughing andunravelling occurs in Units C and D of the ore shale due totime dependent weathering, resulting in the formation of anoverhang brow at the Unit E contact and a ledge at the Unit Bcontact. This leads to localized unstable conditions when thebench is taken. Regional stability Analysis   To establish an idea of the stress levels in the proposedMOCB mining and surrounding areas as a result of earliermining, a mine scale model was set up using FLAC3D(ITASCA, 2004). This indicated that loading on the proposedMOCB areas is not ever more than 40 MPa. In addition, stressvariation within the trial mining block as extraction of thestopes progressed was simulated using a 2-D finite elementcode PHASE2. More complex three-dimensional modelling isplanned for the future as more insight is gained. Model geometry Figure 3 shows the model geometry for the PHASE 2analysis. Model results  The variation of stress as mining progresses across fivestopes is indicated in Figure 4. The stress levels are about 20 MPa at the over-cut raise development stage. Five over-cut raises can be mined at the same time without significantrise in the stress level. The stress level rises to between 35and 45 MPa in the 4 m pillars (P1 to P4) separating the 10 mwide over-cuts and stabilizes at about 30 MPa after thewhole block has been mined and backfilled. The stress levelin the10 m wide ramp pillar (RP) is constant at about 22 MPathroughout the mining stages. As expected, benching of theover-cuts does not increase stress level in the pillars. Thepillar strength is, however, expected to reduce as the w/hratio reduces. This is analysed in detail in the next section. Pillar stability and performance Analysis   To assess the pillar stability and performance in detailanother two-dimensional model using a displacement discon-tinuity boundary element code (Malan 2004) was set-up. Model geometry  The sequence of mining modelled was as follows (Figure 5)   Step 1  —over-cut raise stage—4.0 m wide and 4.0 mhigh excavations at 14 m centres, to define a 10 mwide 4 m high rib pillar/ rectangular pillar (strikelength >/=4 times apparent dip span)   Step 2  —down-dip over-cut raise slyped to full over-cutwidth—10 m wide and 4.0 high opening down-dip and4 m wide by 4 m high openings up-dip to define a 4 mwide and 4 m high rib/rectangular pillar   Step 3  —both down-dip and up-dip over-cut raisesslyped to full stope width—10 m wide excavations bothdown-dip and up-dip leaving a 4 m wide and 4 m highrib/rectangular pillar with 10 m wide excavations oneither side  s 610 OCTOBER 2005 VOLUME 105 NON-REFEREED PAPER The Journal of The South African Institute of Mining and Metallurgy  Figure 2—POTVIN stability graph Table IV  Unsupported span limits Rockmass StabilityHydraulicUnsupported dip spanconditionnumberradius10 m wide12 m Wideover-cutover-cut Poor0.31.6< 10< 10Fair to good3.1415–4010–24Very good368.3>100>100 Table III  Stability numbers Rockmass conditionParametersStability numberClassQ-ratingABC Poor1.10.700.2020.3Fair to good4.60.750.3033.1Very good300.800.30536
Similar documents
We Need Your Support
Thank you for visiting our website and your interest in our free products and services. We are nonprofit website to share and download documents. To the running of this website, we need your help to support us.

Thanks to everyone for your continued support.

No, Thanks