24-MMP-A2 Underground Mining Methods and Design · May 2014
Nivaar worked solution (AI-drafted; not reviewed by a licensed engineer)
EGBC National Exam — Mining and Mineral Processing Engineering, 09-MMP-A2 Underground Mining Methods and Design, 2014-May. 3 hours duration, closed book; only a Casio or Sharp approved calculator permitted. Question 1 is compulsory (40 marks, all seven parts 1.1–1.7); a candidate then selects FOUR of Questions 2–7 (each worth 15 marks).
Reference texts: Hartman & Mutmansky (eds.), SME Mining Engineering Handbook, 3rd ed. (underground mining methods, mine ventilation, shaft hoisting design, headframes, backfill practice, mine cost estimation — the primary reference throughout this paper); Hustrulid & Bullock, Underground Mining Methods: Engineering Fundamentals and International Case Studies (narrow-vein longitudinal-retreat/Avoca-family stoping, cut-and-fill variants); BC Ministry of Energy, Mines and Low Carbon Innovation, Health, Safety and Reclamation Code for Mines in British Columbia (Canadian regulatory context for hoisting-rope safety factors and shaft ventilation); Camm, T.W. (1991), Simplified Cost Models for Prefeasibility Mineral Evaluations, U.S. Bureau of Mines IC 9298 (source of the Question 4 parametric cost models).
Question text not reproduced: the examination questions are © Engineers and Geoscientists BC. Open the official past paper (linked at the top of this page) to read the question, then follow the worked solution below.
2.1.1 Ore handling system. The shaft is collared and sunk in the footwall, offset laterally from the vein by enough distance to remain outside the eventual stress-relaxation/subsidence zone of the stoped-out orebody and to avoid sterilising ore — a common rule of thumb is an offset of at least one shaft-pillar radius, refined by the geotechnical assessment in 2.1.3. Five main haulage levels are developed at 500, 600, 700, 800 and 900 m, connected to the shaft by crosscuts, with two sub-levels between each pair of main levels (matching the given spacing) accessed from the main level by ramp or internal raise, from which stope drawpoints are developed. Broken ore is mucked at each sub-level by LHD to a transfer (ore-pass) raise that drops the ore to the main haulage level below; on the haulage level, trackless haulage (trucks, or LHD tramming for shorter distances) carries ore along the transport drift and crosscut to a shaft-bottom loading pocket, typically via a primary crusher station sited near the shaft to size the ore for efficient, consistent skip loading and to reduce hoisting-cycle variability. Mobile equipment: LHDs at every sub-level for stope mucking to the ore-pass; haul trucks (or a second stage of LHD tramming) from ore-pass draw stations to the crusher/loading pocket; and a skip-hoisting system (of the type designed in Question 6) from the loading pocket to surface.
2.1.2 Mine services. Clean water is piped from a surface tank down a dedicated range in the shaft's service compartment, with take-offs at each main level feeding drilling and dust-suppression water to the sub-levels. Electrical power is fed from a surface substation down an armoured trailing/shaft cable to underground substations at alternating main levels, stepped down locally to utilisation voltage for face equipment; the shaft cable is sized for the ultimate 3000 t/d, three-shift production load. Compressed air is reticulated from surface compressors through a shaft air range with level take-offs, sized for the coincident peak demand of drilling/loading equipment across the levels being actively developed at any time; a booster station may be added at depth once static-head losses in a 400+ m column become significant. Waste water (mine drainage) collects by gravity in sumps at the lowest active level (900 m) and is pumped in stages — a main pump station at 900 m lifting to an intermediate sump, then to surface — sized for natural groundwater inflow plus service water. Communications (leaky-feeder radio and hardline/data cable) run down the shaft alongside power, with repeaters/nodes at each level tying back to the surface control room for voice, tracking and emergency signalling.
2.1.3 Design factors and sketches. Key factors: (i) shaft position in the footwall, set back far enough to avoid ore sterilisation and stope-induced ground movement, but close enough to keep crosscut development short; (ii) geotechnical conditions along the shaft route (avoiding major structures, weak zones or water-bearing horizons); (iii) level spacing (the given 100 m main-level/2-sublevel pattern) balancing lateral development cost against stope size and ventilation-circuit segmentation; (iv) ventilation-circuit integration, keeping service raises and waste-water infrastructure out of return-air airways; (v) ore/waste segregation at the transfer points to protect mill-feed grade; and (vi) sizing every service (power, air, water, hoisting) for the full 3000 t/d × 3-shift, 6-year mine life rather than an initial ramp-up rate, since retrofitting shaft-compartment infrastructure after sinking is disproportionately expensive. The accompanying figure is the hand-drawn-equivalent layout: shaft in the footwall, five main levels at the given depths, two dashed sub-levels between each, all levels crosscutting to the dipping vein.
