V A 9 r. ** a* ** o, •Tr»« a ^» '• ^ .v^ra- ^ ;£K&»*.. ^ :<^»: "ov* ?a&*~ '++& «5^ .0*' *o, *,TTT*' A ^6* ■ U A* .!- <>. *..•' \^ ■A? .»W«w* > f" e^^ -^H^« A^"x. • -*. 4 ***< A * ^."^^'a^ ^ :^£^ \>^ /jOk^ **^ .*^BaC0Sr- X/ ••; *\ J*% sP««» * aV <* • 6 T 'X **v ,o. ++< *6 4 A * *^r$ Sa- "*o^ o«s la- ^ * r \ *+ 0# * aV^ 33 ELEVATION Not t0 scale Figure 4.-Discharge vehicle. ELEVATION Figure 5. -Bridge conveyor. Not to scale SURFACE TESTING JEFFREY MINING MACHINERY TESTING During the 1977-79 period, JMMD conducted a series of tests on the MUCH system at its facility. The objec- tives of these tests were to verify the tracking-retracking ability of the system in 90° room-and-pillar configuration in an aboveground environment and ascertain the convey- ing capabilities of the system. These tests revealed that each vehicle would overshoot the path of preceding vehicle and the MUCH system train would not track parallel to the section belt. The over- shooting problem was solved by increasing the steering bar length by 2 in. The problem of the MUCH train not tracking parallel to the section belt was corrected with addition of hydraulic power for steering the discharge vehi- cle. The hydraulic power pack was added to the bridge conveyor, and the control valve and steering cylinder were added to the discharge vehicle. Other problems, such as jackknifing, were not solved although many modifications and/or changes were made in the system. The results of 1977-79 testing thus remained inconclusive. In 1980, JMMD was asked to conduct additional testing on the MUCH system. The objectives of additional testing were to find out the MUCH system limits, its ability to operate and stop on inclines, and how it would track through S-turns on various slopes, over the rolls, and over loose and mud bottom. These tests were conducted at the Ohio Transportation Research Center (OTRC). A three-entry mine plan was laid out at the OTRC. The simulated walls of entries and crosscuts were con- structed with snow fence. The system would not track- retrack in the same manner at the OTRC as it had at the Jeffrey facility. It required modifications and changes for steering and wheel alignment, as well as addition of turn- buckles to all vehicles and brakes on the lead and dis- charge vehicles. After these modifications and changes were made, there was an improvement in track-retracking and conveyor discharge performance. In 1982, additional design changes were made prior to sending the system to the Bureau's test facility at Bruce- ton, PA. These changes included installation of a chain between hopper and conveyor to limit travel between the adjacent vehicles, installation of rubber belting on hopper sideboards to stop spillage, and relocation of the conveyor speed switch. METF TEST PROGRAM OVERVIEW The MUCH system, consisting of 12 vehicles, bridge conveyor, and numerous spare parts, was received at the Bureau's METF in July 1983 for assembly and surface testing. Upon completion of assembly, all functions and safety devices were checked out and made operational; operator familiarization and training was undertaken prior to starting the test program. Surface tests were conducted at the METF to verify and evaluate the performance of the MUCH system. Tests were divided into sequences to evaluate a particular sub- system or machine function. Modifications were made to the MUCH system to correct deficiencies noted during surface testing. A description of each test sequence is given in the following sections. TRAM AND MANEUVERABILITY TRIALS Tests were conducted in the METF equipment maneu- verability trial area (EMTA) to determine the tramming capability of the system in a simulated mine environment, to define and correct any observed tramming or steering problems, and to demonstrate the reliability of the overall tram system. Early in the surface test program, it became obvious that tramming and maneuvering the 12-vehicle system was no simple matter, especially within the confines of the EMTA and with inexperienced operators. The system was trammed into a continuous loop numerous times for haul- age system trials, demonstrations, and noise level tests, and was also trammed in and through the simulated mine workings of the EMTA. Figure 6 illustrates a number of operational procedures and maneuvering sequences that were undertaken early in the tramming trials. Initially, these maneuvers were diffi- cult to perform successfully. Vehicles were contacting the ribs at corners as well as midrib. After operators per- formed a number of maneuvers, they were better able to gauge how an entry or crosscut must be entered, how tight a radius to turn, when to begin straightening the lead vehi- cle in a turn, how much rib clearance was needed, etc. Figures 6B through 6F illustrate typical maneuvers per- formed to place change the system from a far-right entry to the left crosscut in a simulated three-entry section. Figures 6F through 6H illustrate the return of the system to the far-right entry. In the step shown in figure 6£>, it was necessary to reposition the outby end of the system to provide sufficient room to accommodate the limited turn- ing ability of the discharge vehicle. During these early tramming trials, more problems were experienced with the system making rib contact near the center of the pillar (area 1, fig. 6H) than at the pillar corners (area 2, fig. 6H). At the pillar corners, when making a 90° turn, the system tended to move away from the corner, but at midpillar, the system worked toward the rib and sometimes made contact. In general, the overall tracking of the system seemed inconsistent, but the track- ing seemed better when turning sharp 90° turns than when making more gentle, larger radius turns. Figure 67 shows the system in a gentle S-curve being trammed through the EMTA and out the east equipment door. This maneuver proved to be one of the most diffi- cult encountered during the early tramming trials. Three pillar corners and the exit door frame were contacted numerous times while the system was being trammed out of the building. It was much more difficult to move the system away from the rib or corner when the train was relatively straight than when the train was in a tight turn. This was expected, because the force available to pull a vehicle away from a rib or corner is proportional to the sine of the angle between the jammed and adjacent vehicle, and dependent upon the radius of curvature of the train. The typical method used to unjam vehicles from contact with a pillar was only marginally effective when the train was in a gentle curve. The method consists of shutting off outby vehicles and tramming the remaining vehicles inby, then pulling vehicles away from the rib. Likewise, unjam- ming vehicles from contact with a pillar corner by remov- ing tram power from outby vehicles and tramming remain- ing vehicles in the outby direction to push jammed vehicles away from the rib was ineffective when the train was in a gentle curve. These methods did work quite well, how- ever, when the system was in a sharp turn. A coal ramp approximately 2.5 ft high (fig. 6/) was built to observe system performance as vehicles trammed over it. The lead vehicle and first two intermediate vehicles were successfully trammed over the coal ramp and down the other side despite deep wheel penetration. No serious mechanical interferences were observed. Steering bars provided the necessary degrees of freedom in roll, pitch, and yaw axes. As the overall test program continued, a vehicle-to-vehicle interference problem that led to major vehicle frame modifications became obvious. This problem and subsequent modifications are described in the "System Modifications" section of this report. As a result of these initial tramming trials, the steering- tracking system was examined to better in defining the theoretical capabilities of the system and correct any mechanical problems which might hinder tracking ability. Steering System A physical layout of the mechanical steering system (fig. 7) was constructed that reflected the linkage geometry of one steering axle and drawbar. This model led to a better understanding of the steering system and helped in defining the theoretical limits of the system. Figure 8 shows an overall layout of the system making two turns of different radii. During tramming of the system, it had been observed that the tracking seemed to be better when making tight turns than when making more gradual turns. The tracking layout drawing supports this observation. It can be seen that when the system is making a 12-ft-radius turn, the center of the turn is coincident for all vehicles making the turn. A 43-ft turning radius is also shown in figure 8. It can be seen that the center of the turn wanders as the turn is entered by successive vehicles. It moves approximately 14 in with each new vehicle entering the turn. An en- larged view (fig. 9) shows the tracking error of each vehi- cle as the vehicles pass the same reference point. A track- ing error of 5.25 in per vehicle occurs when turning at a 6° drawbar angle, which yields a 43-ft turning radius. Once the tracking was defined from a theoretical stand- point, it was obvious that the system was not tracking as well as could be expected; therefore, the mechanical steer- ing components of each vehicle were examined. The steer- ing system freeplay was measured with a dial indicator on the end of each steering bar where it attached to the tie rods. Freeplay ranged from 0.03 to 0.050 in. The causes of excessive freeplay were found to be two loose tie rods, which were tightened, and 26 loose steering plates on 14 axles. The steering plates are bolted to the steering knuckles by four bolts on each plate. Once the bolts loosen, the plate is free to rotate as much as the clearance in the boltholes will allow, thus allowing excessive freeplay in the system. All loose steering plate bolts were torqued to specifications. *48'-»H20W 20' "T 48' n: \ r Lead vehicle Continuous loop haulage demonstration ]oD D Tram into entry «3 ..J D D □ g- — o D Tram ahead to provide room to back up and steer discharge car • ( » D D Tram out by f V 1 n i l \ / Tram outby to place change □ Dl ^ !<*&- H D D Begin simulated place change JJO D ..J Complete place change into new entry J % D D 1 / Tram over coal pile I I Figure 6. -MUCH system operational maneuvers in equipment maneuverability trial area. Drawbar Figure 7.-Mechanical steering linkage geometry. 10 Figure 8. -Tracking layout, 12- and 43-ft turning radii. The mechanical steering system on each vehicle was properly aligned per instructions in the operator's manual. Adjacent vehicles were trammed into a straight line and the drawbars were aligned along the axis of the vehicles. The tie rods were then adjusted so that each wheel was parallel to the drawbar center line. The elimination of excessive freeplay in the steering system and the alignment of the individual wheels had a positive effect on the system tracking, but some problems were still evident, especially when tramming in the outby direction. Tramming outby was difficult because of the severely limited steering range of the discharge vehicle and the tendency of the vehicles in the trailing half of the system to track poorly and jackknife. Modifications were made to the discharge vehicle steering system to increase the available steering angles and improve component location. The modifications are described in the "System Modifications" section of this report. The tendency of the trailing half of the system to track poorly and jackknife while tramming in the outby direction was investigated and the problem was corrected. The tram system is designed so that the electrical power to the tram motors in the last two trailing vehicles is automatically cut off during tramming to provide tension in the train. When the tram control in the lead vehicle is actuated, all tram motors receive power, but after approximately 1.1 s, the motors in the two trailing vehicles are automatically pow- ered down to create drag on the vehicle train. Upon investigation, a malfunctioning time-delay relay was found in the lead vehicle outby tram circuit that did not power the motor down after 1.1 s of operation. The unit was disassembled and repaired. Subsequently, track- ing in the outby direction improved and jackknifing when tramming outby was reduced, because additional drag was provided by the lead vehicle. Tracking-Retracking Tests Objective The objective of the tracking-retracking tests was to determine the ability of the MUCH system to successfully tram in both the inby and outby directions within the con- straints of the EMTA simulated workings. 11 Direction of travel Approx A _ B tracking misalignment =5/4 in Figure 9.-Tracking capability, 43-ft turning radius. Procedure The EMTA was utilized to simulate an underground working area with 48- by 48-ft pillars and 20-ft-wide entries and crosscuts, as shown in figure 10. Ten data stations were located on the test course. Each data station con- sisted of a heavy string pulled taut across the entry approx- imately 5 ft above the floor. A 20-ft-long by 1-ft-wide strip of heavy kraft paper was suspended from each string. Uncapped felt-tip markers of assorted colors were attached to each of the 12 MUCH system vehicles, one per vehicle, at the same location on each vehicle. As the system was trammed through the test course, the markers were drawn across the kraft paper at each data station to permanently record the relative position of each vehicle at each station. To conduct the tests, the entire MUCH system was trammed through the course of 10 data stations four times, twice in the inby direction and twice in the outby direc- tion. After each traverse, the positions of both the lead and discharge vehicle, as indicated by the associated felt- tip mark on the kraft paper at each data station, were measured relative to both the left and right ribs. While tramming in the inby direction, the system was controlled by the steering of the lead vehicle. In the outby tram direction, the discharge vehicle steering controlled the system. 12 KEY ~ j Simulated ribs System tracking variance 10 Data stations Figure 10.-Trial A (top) and trial B (bottom) inby. 13 Results and Discussion The results of the tracking-retracking tests are given in table 1 and are shown in figures 10 and 11. Table 1 pre- sents the analytical data acquired during the trials. Fig- ures 10 and 11 present layout drawings of the test course that show the dimensions of the course, data station loca- tions, direction of system travel, path of the operator- controlled vehicle through the course, and tracking error (variance) of the system through the course. In table 1, the operating width column shows the max- imum width required by the system at the associated data station. This width is equal to the width of the operator- controlled vehicle (6 ft 8 in) plus the maximum amount of tracking error (variance) shown by any other vehicle at the data station. The maximum variance column shows the maximum variance observed at each data station and whether the variance was to the left or the right of the position of the operator-controlled vehicle. The rib dis- tance column shows the positioning of the required operat- ing width of the system in reference to the left and right ribs at the data station. The left and right designations are relative to the operator position, which is always with the first vehicle facing the direction of travel. During the tramming trials, the MUCH system tended to drift toward the inside radius of the turns. This tight- ening in the curves is probably due to the tramming resis- tance imposed by the automatic braking of the last two cars on the trailing end of the system. This braking keeps the system in tension, which keeps individual vehicles from jackknifing, but also tends to pull the vehicles toward the inside of a curve. The maximum variance that occurred while tramming in the inby direction was 68-in, which occurred at data station 5 during trial A (table 1, fig. 10). Tramming in the outby direction was more difficult than tramming inby because of the more limited steering capacity of the discharge vehicle. During both trials, rib contact was made at the corner where data stations 5 and 6 intersect (fig. 11) when tramming in the outby direction. This corner was also the point of maximum system vari- ance, 94- and 101-in, during trials A and B, respectively. During trial A, intermediate vehicle 4 contacted the cor- ner. In trial B, as shown in figure 11, the system was trammed closer to the left rib prior to starting the turn to the right, and the turn was initiated sooner to create a smoother flowing curve. During this trial, only the last vehicle intersected the rib, although the overall variance was greater than in trial A. Because of the lightweight fiberglass panels in the EMTA, the rib contact during tramming was a problem that prevented further tramming until the system was moved away from the rib. In a similar underground situa- tion where rib contact would cause no damage, the system could have continued tramming while being guided by the rib. While rib contact is not desirable, it would not be as much of a problem underground as in the EMTA. In the author's opinion, the system tracked and retracked with enough consistency to operate satisfactorily in an under- ground mine of similar dimensions. Simulated Production Cycle Trial Objective The objective of the simulated production cycle trial was to evaluate the tramming and tracking ability of the MUCH system during a simulated face production cycle. Procedure The EMTA was utilized to simulate an underground working area with 48- by 48-ft pillars, 20-ft-wide entries, and 20-ft-wide, 90° crosscuts. A National Mine Service shuttle car was utilized to simulate a continuous miner at a production face. The simulated production face was in the center entry of the three-entry simulated workings. The discharge vehicle of the MUCH system was in the left-hand entry; the 10 intermediate vehicles ranged from the discharge vehicle, through a 90° right crosscut to the center entry, through 90° left-hand turn to the lead vehicle, which was positioned directly behind the simulated miner at the face (fig. 12). A continuous miner cut cycle was simulated by advanc- ing the simulated miner and MUCH system 2 ft to simu- late a sump and shear cycle, retreating 4 ft to prepare to cut the cusp and clean up, then advancing 6 ft to cut the cusp, clean up, and sump and shear. This cycle was re- peated 10 times to simulate a 20-ft face advance. Upon the completion of this cut cycle, the simulated miner and MUCH system were backed up approximately 30 ft, the simulated miner was repositioned to the left-hand side of the face, and the cut cycle was repeated on the second lift. The overall functioning of the tram system, system alignment, and tire tracks were observed during the simu- lated cutting cycles. Upon the completion of the simulated cutting cycles, the MUCH system was trammed to simulate a place change. The MUCH was trammed outby, down the mid- dle entry, through the crosscut, and back into the left- hand entry. The system was then trammed straight ahead in the left-hand entry, past the crosscut and inby into the next 90° crosscut, and to the right to simulate a place change. The time required to complete with place change was measured. Results and Discussion The MUCH system trammed without problem and tracked fairly well during these trials. No rib contact, jack- knifing, or tramming delays occurred during tests. The tracking of the system during the simulated cut cycles was good. As was evidenced in the previous tracking-retrack- ing tests, the system tends to wander toward the inside radius of the 90° turns. This tendency was also observed during these trials but to a lesser degree. The maximum tracking variance observed during these trials is shown in figure 13. This series of tire tracks shows the tracking variance at the 90° turn located at the intersection of the crosscut and middle entry. This 14 "footprint" is the accumulation of tire tracks created while the system was trammed inby to the face, operated through a 20-ft cutting advance, a lift change, a second 20- ft cut advance, and then trammed outby during the place change. The total width of the track is 35 in. Therefore, the MUCH system required a total working width of 35 in plus the width of the lead vehicle to operate without rib contact. The total required operational width would be 9 ft 7 in at that location. KEY I | Simulated ribs MM System tracking variance I -10 Data stations Figure 1 1 -Trial A (top) and trial B (bottom) outby. 15 Table 1.-MUCH tracking-retracking trials Data Operating width Max variance, in Rib distance station Left Right Left Riqht ft in ft in ft in Trial A, in by: 1 2 3 4 5 6 7 8 9 10 Trial B, inby: 1 2 3 4 5 6 7 8 9 10 Trial A, outby: 10 9 8 7 6 5 4 3 2 1 Trial B, outby: 10 9 8 7 6 5 4 3 2 1 *Made contact with rib. 8 7 10 10 12 10 8 8 10 9 9 9 7 8 9 9 7 8 7 9 9 8 9 8 12 13 8 8 8 8 9 8 8 10 14 11 9 8 8 11 8 10 6 7 4 2 4 4 8 5 7 7 1 10 10 2 6 11 7 10 4 9 1 10 8 6 9 8 7 2 8 8 4 3 3 10 11 2 24 14 46 47 68 40 8 33 35 11 17 36 36 44 34 26 33 32 44 26 32 35 61 10 20 44 36 6 22 15 28 14 33 86 94 32 30 55 101 70 39 34 6 5 1 1 2 3 6 8 7 6 4 4 6 6 4 7 6 7 6 6 1 1 4 10 7 6 9 6 4 1 2 4 6 5 8 10 9 5 1 4 10 9 9 3 11 8 10 11 8 5 4 9 9 3 5 4 5 1 7 10 2 11 2 7 11 3 11 5 6 10 7 8 9 2 7 5 3 5 6 7 7 5 6 5 2 2 3 5 6 6 5 5 2 6 4 5 4 9 9 5 1 1 4 7 10 9 9 6 2 1 5 6 4 9 8 5 1 2 10 8 5 11 5 7 5 5 7 1 9 7 4 7 10 1 J '0 9 7 5 4 6 :? *o 7 7 8 11 Figure 12.-Lead vehicle behind shuttle car. 16 Conveyor Noise-Level Survey Objective Figure 13.-Tracking variance. The simulated place change that involved tramming the system outby from the middle entry, through a crosscut into the left-hand entry, then inby in the left-hand entry to the next crosscut and turning into the crosscut, covered a total tram distance of 245 ft. The outby tramming for the place change covered a distance of 155 ft and required a total time of 1 min 30 s, there was a pause of 10 s while the lead vehicle operator communicated with the discharge vehicle operator before starting to tram inby to complete the place change. The inby tramming covered a distance of 130 ft and required 1 min 40 s. The complete place change required a total of 3 min 20 s to complete and no problems occurred during the trial. CONVEYOR SYSTEM TESTS The main purpose of the MUCH system is to continu- ously convey coal cut by a continuous miner at the face to the section panel belt. Therefore, a large portion of the MUCH test program was devoted to testing, evaluation, repairs, and modification of the conveyor system. A conveyor noise level survey was conducted to deter- mine noise exposure experienced by operators or miners working close to the MUCH system. Procedure The system was set up in a continuous loop behind the EMTA. The train was configured in a circle with the discharge vehicle dumping into the lead vehicle hopper while operating in the conveyor mode only. Tests were performed with the conveyors empty and with varying amounts of coal in the conveyors. Coal was a mixture of 2.5- by 2-in and 2- by 1.5-in sizes. All tests, except ambi- ent noise measurement, were performed with the convey- ing system in operation. Noise-level measurements were obtained in the center of the vehicle circle, around the outer perimeter of the vehicle train, and in the operator compartment. Attention was focused on the operator compartment of the lead vehicle. A Bruel and Kjaer type 2205 handheld sound level meter (SLM) with a Bruel and Kjaer type 4117 piezo- electric microphone was used to perform the sound level measurements. Sound pressure level (SPL) was measured using the A-weighted network. This filter network patterns its response after the human ear. SPL meter response was in the slow mode. All measurements were taken with the meter held away from the body to minimize the effects of noise reflection from the body. It should be noted, however, that perform- ing a noise survey in an enclosed building, particularly with a large number of hard surfaces reverberating or reflecting noise, may cause SPL measurement errors. Larger objects with physical dimensions similar to the wavelength of the sound being measured are most likely to reflect noise and to be sources of error. Considering these limitations, the test area was adequate to perform such a general noise survey. Results and Discussion The noise level test configuration is shown in figure 14 and the survey results are given in table 2. The three tests are discussed in the following sections. Test 1-Conveyor Off, Ambient Noise Level Measurement The measured ambient noise level of 50 to 51 dBA in the test area introduced no error into the operating noise level measured. Ambient noise level greater than 10 dBA below the measured SPL will introduce no significant error in the data; therefore, no compensation for ambient noise level was made. 17 Table 2.-Conveyor noise test results (sound pressure levels) Configuration Conveyor off, ambient noise level Conveyor operating, no coal in conveyor: Reading from center of conveyor loop (sound level meter rotated 360°) Pass around (except at lead and discharge vehicle junction) Junction of lead and discharge vehicle Operator's ear level in lead vehicle Conveyor operating with coal, level at lead vehicle operator's ear: Approximately 3.7 st/min Approximately 7.4 st/min Approximately 11.1 st/min dBA 50 - 51 98.3 101 -102 103 -104 102 98 97 96.5 - 97.5 KEY Sound level meter EMTA Equipment maneuverability trial area EMTA Discharge vehicle Lead vehicle Operator compartment Figure 14.-Conveyor noise-level test configuration. 18 Test 2— Conveyor System On, No Coal a. The SPL was measured in a 360° arc from a position in the center of the conveyor loop (fig. 14). The SLM was held away from the technician's body to minimize error, 4 to 5 ft above the ground and pointing toward the outside of the vehicle train circle. The measured SPL was 98.3 dBA. b. A pass-by noise measurement was made by walking parallel to the train on the outside perimeter, approxi- mately 2 ft from the vehicles. The SLM was held 4 to 5 ft above the floor and away from technician's body. The axis of the SLM was held parallel to the tangent of the MUCH circle, approximately 2 ft away from the vehicles. The SLM was held parallel to the vehicles to maintain geo- metric uniformity and to facilitate meter reading. The pass-by noise measurement is most closely associated with the noise level that a nearby miner would experience when in the same entry with the system. The measured SPL ranged from 101 to 102 dBA. c. The SPL measured at the lead-discharge vehicle junction was 103 to 104 dBA in the pass-by mode as in test 2b. It was slightly more noisy than the junction of any other two vehicles. d. The SPL in the operator's compartment at the operator's ear level was measured to determine noise level exposure of the lead vehicle operator. The operator com- partment, at 102 dBA, was very noisy. Moving the micro- phone around the vicinity of the operator's ear showed minimal change in the readings, and thus increased the confidence in the SPL measurement. Accurate meter positioning was, therefore, found unnecessary. In addition to customary hearing protection, a head- set for communicating with the discharge vehicle opera- tor would significantly attenuate the noise level to the operator. Test 3-Conveyor System On, With Coal A front-end loader was used to meter approximately 3 yd 3 of coal onto the MUCH system. The SPL at opera- tor's ear level was measured while the system was con- veying coal at approximately 3.7 st/min. The introduction of coal into the conveying system, even in small amounts, significantly reduced the noise level by damping the con- veyor structure and absorbing acoustic energy. Noise level at the operator's ear was reduced by 4 dBA to 98 dBA. The test was repeated, this time with 6 yd 3 of coal in the conveying system (7.4-st/min haulage rate). A 1-dBA noise level reduction was achieved by this doubling of the amount of coal on the conveyor (from 98 to 97 dBA). The test was again repeated with 9 yd 3 of coal in the conveyors, an 11.1-st/min haulage rate. No further reduc- tion in noise level over the 6-yd test was achieved by the addition of 50 pet more coal (SPL, 96.5-97.5 dBA). It appeared that adding coal to the conveying system signifi- cantly reduced operator noise exposure, but further reduc- tion in noise level was minimal beyond the 3.7-st/min haulage rate. The noise levels measured indicate that some form of hearing protection will be required for the MUCH system operator to maintain an 8-h noise exposure of 90 dBA or below. Conveyor Speed Test Objective A conveyor speed test was conducted to determine con- veyor chain speed and motor rotational speed under both loaded and unloaded conditions and to determine the transport speed of discrete particles at various loading rates. Procedure A Micronta 63-5009 digital stopwatch was used to mea- sure time required for the conveyor chain in each vehicle to make a complete cycle. When the system was loaded with coal, the conveyor chains were not visible; therefore, a General Radio 1531AB stroboscope was used to measure tailshaft rotational speed. After establishing a relationship between chain speed and tailshaft rotational speed, the chain speed was measured with a stroboscope. A stop- watch was also used to measure the transport speed of individual chunks of coal. Results and Discussion The conveyor surface speed was measured when empty and when loading 2 and 7 st/min (table 3). At st/min (empty), the time required for the conveyor chain to make one complete pass through the vehicle was measured with a stopwatch. Three trials were conducted for each vehicle in the empty condition. The conveyor chain lengths were 45 ft 5 in for the lead vehicle and 42 ft 5 in for all other vehicles. Dividing chain length by time yielded the average conveyor speed for each vehicle, which ranged from 270.7 to 273.7 ft/min, an average of 272.2 ft/min. The 272.2-ft/min average agrees within 2.7 pet of the published figure of 280 ft/min in the work cited in footnote 4. Conveyor tailshaft speed was also measured to deter- mine the speed relationship between the conveyor and tailshaft. Tailshaft rotational speed averaged 266.0 r/min under empty conditions. Thereafter, when coal was in the system, the conveyor speed was measured via tailshaft rotational speed. The ratio of conveyor speed to tailshaft rotational speed (K r/min ) was 1.025. Multiplying tailshaft rotational speed by K r/min gives the equivalent chain speed. At a conveying rate of 2 st/min, conveyor chain speed averaged 268.3 ft/min. At 7 st/min, average chain speed was 270.9 ft/min. From to 7 st/min, average conveyor speeds were within 1.6 pet of one another. At 2 st/min. 19 average conveyor speed was 2.6 ft/min slower than at 7 st/min. This could be attributed to the small sample taken at 2 st/min (four vehicles) and the possibility of fines buildup on the conveyor deck during the trial. Transport speed of individual particles was measured at various loading rates. Several chunks of coal and coalcrete (from 2 to 6 in) were timed as they passed through the system (table 4). With no coal in the system, the average transport speed was 253.5 ft/min. This figure is about 7 pet less than the average conveyor speed of 272.7 ft/min, reflecting the time required for the conveyor chain to pick up coal as it cascaded onto the next outby vehicle. With the system conveying about 2 st/min, transport speed was about 3.2 ft/min less, or 250.3 ft/min. At 4 st/min, transport speed was 234.0 ft/min, but only one trial was performed. Generally, transport speed will decrease somewhat at higher loading rates. From observation, it can be seen that some coal tends to ride over the chain flights as load- ing rate increases, thus tending to decrease the transport rate. It was also observed that as coal became finer from repeated handling, a greater percentage of it tended to ride up over chain flights, further reducing average trans- port speed. With relatively large pieces (2 to 6 in), how- ever, no correlation could be seen between particle size and transport speed. Another way of interpreting transport speed is to relate it to conveyor speed, providing a measure of conveying efficiency. At st/min, transport speed is 253.5 ft/min and conveyor speed is 272.7 ft/min. Transport speed (or conveying efficiency) is 93 pet of the conveyor speed. Con- versely, slippage would only be 7 pet. A 100-pct efficient system would transport coal at the same speed as the con- veyor travels. At 2 st/min the effective conveyor transport efficiency is 93.3 pet, and at 4 st/min it is 87.3 pet. Table 3.-Conveyor speeds Time per convevor evele, s Tailshaft Av speed, conveyor speed, 1 Conveyor speed Vehicle Trial Trial Trial to tailshaft ratio 1 2 3 r/min ft/min ( K r/min)- r / min HAULAGE RATE: st/min (SYSTEM EMPTY)-3/28/84 Lead 10.04 9.88 10.02 266 273.1 1.027 1 9.49 9.31 9.39 265 270.7 1.022 2 9.34 9.54 9.30 266 271.0 1.019 3 9.30 9.33 9.29 266 273.4 1.028 4 9.27 9.34 9.35 266 273.1 1.027 5 9.39 9.31 9.26 266 273.1 1.027 6 9.27 9.30 9.32 267 273.7 1.025 7 9.43 9.37 9.28 266 271.9 1.022 8 9.31 9.28 9.30 266 273.7 1.029 9 9.30 9.34 9.25 265 273.7 1.033 10 9.31 9.35 9.35 266 272.5 1.024 9.26 9.32 9.42 267 272.8 1.022 Average NAp NAp NAp 266.0 272.7 1.025 HAULAGE RATE: 2 st/min-3/14/84 8 NAp NAp NAp 264 270.6 NAp 9 NAp NAp NAp 257 263.4 NAp 10 NAp NAp NAp 263 269.6 NAp Discharge NAp NAp NAp 263 269.6 NAp Average NAp NAp NAp 261.8 268.3 NAp HAULAGE RATE: 7 st/min-3/30/84 NAp NAp NAp 262 268.6 NAp 1 NAp NAp NAp 264 270.6 NAp 2 NAp NAp NAp 264 270.6 NAp 3 NAp NAp NAp 264 270.6 NAp 4 NAp NAp NAp 264 270.6 NAp 5 NAp NAp NAp 263 269.6 NAp 6 NAp NAp NAp 263 269.6 NAp 7 NAp NAp NAp 268 274.7 NAp 8 NAp NAp NAp 266 272.7 NAp 9 NAp NAp NAp 263 269.6 NAp 10 NAp NAp NAp 266 272.2 NAp Discharge NAp NAp NAp 264 270.6 NAp Av NAp NAp NAp 264.3 270.9 NAp NAp Not applicable. 'Calculated from total length of conveyor chain: lead vehicle-45 ft 5 in, all other vehicles-42 ft 5 in. 20 Table 4.-Particle transport speed Time. 1 s Particle diam, in Trial Trial Trial 1 2 3 0-st/min haulage: 4 53.40 53.44 54.09 4 54.69 NT NT 6 53.87 54.55 54.89 2-st/min haulage: 2 57.10 NT NT 5 53.02 54.99 NT 6 54.35 NT NT 4-st/min haulage: 4 58.66 NT NJ_ NT No trial. transport time from lead vehicle hopper to end of discharge vehicle. Av Transport speed, ft/min Av for all sizes at loading rate 255 250. 252. i} 240.4 254.1 252.5 234.0 253.5 250.3 234.0 Conveyor Time Sequencing Objective A conveyor time sequencing test was conducted to measure the startup sequence time required to start the conveyor system. Procedure Time required to power up all conveyors sequentially was recorded under loaded and unloaded conditions. During testing, with haulage rate established and coal evenly distributed throughout the MUCH system, the time required to restart the system was measured by the lead vehicle operator using a digital stopwatch. Time was measured from the time the conveyor start switch was pushed until the lead vehicle conveyor system started. Results and Discussion The time required to start up the entire train of convey- ors was recorded at haulage rates of 0, 2, 3, and 4 st/min. From data in table 5, it can be seen that the average startup ranged from 1.36 s at no load (0 st/min) to 1.86 s at 4 st/min, generally increasing as the load on the convey- or increased. On March 14, 1984, it took 2.73 s to start up the system at 2 st/min, an apparent anomaly. This was the first date of testing, and solidified coal had been in the idle convey- ors for a period of several days before the testing started. The additional load imposed by this material caused a significantly longer startup time. Coal Conveying Tests Numerous coal conveying tests were conducted during the program and all but the final test were followed by system modifications and improvements in an effort to establish the coal-conveying capability of the system and to achieve a 95-pct system availability while conveying at a rate of 8 st/min over an 8-h shift. March 14 to April 12, 1984 Test Configuration A continuous haulage loop of the MUCH system was formed by having the discharge vehicle convey coal onto the 30-ft Long-Airdox belt structure and into the hopper- feeder-bolter (HFB). (Because of its ability to provide surge capacity, the HFB served as the coal-entry point for new coal being added to the system.) The HFB in turn dumped onto a 50-ft Long-Airdox belt structure, modified to accept the Ramsey Engineering belt weight scale, that loaded coal onto the lead vehicle of the MUCH system, thus completing the loop, as shown in figures 15 and 16. Coal used during the first part of testing was a mixture of 2.5- by 2-in and 2- by 1.5-in coal acquired from the Bureau's Hydraulic Transport Research Facility (HTRF). The second portion of conveyor testing was performed with run-of-mine (ROM) coal from the Bureau's research mine. Table 5. -Conveyor startup time Trial Date Startup time, s Sequenced Av 0-st/min haulage: s ::::::::::::: [ ^^ \ !^ [ 136 2-st/min haulage: 1 "1 J2.83 "1 I ::::::::::::: \ w* \ '?:§ \ 273 4 J I 1.59 J !:::::::::::::} 3 ^ { IS 1 ,„ !:::::::::::::> ***« { !S J 3-st/min haulage: 1 I f 184 1 2 > 4/03/84 < 1.81 > 1.87 3 J I 195 J 4-st/min haulage: 2 ::::::;::::::} v™'** { I:* } 186 'Measured during initial startup of conveyor testing. Solidified coal was present in conveyors, which increased startup time by a significant margin. 21 Figure 1 5.-Conveyor test configuration. MUCH, system Lead vehicle v. Conveyor drive gearbox Thermocouples Explosion- proof box^^ Transducer package HFB Belt Belt scale Belt scale electronics Thermocouple readout Recorder nstrumentation trailer Figure 1 6.-Test configuration and instrumentation system. 22 Instrumentation To more effectively evaluate the conveying system, pertinent electrical and mechanical parameters were mea- sured and recorded during conveying tests. Shown in figure 16 are the layout and location of the instrument systems. The following sections describe parameters mea- sured and instrumentation utilized. Total Conveying System Power Total system power was measured to provide an eval- uation of power requirements of the system, particularly under various operating conditions. True power, in kilowatts, of all 12 vehicles was mea- sured with a Rochester Instrument Systems watt trans- ducer. Current to the watt transducer was supplied by two 400:5 current transducers, one on each of two phases of the incoming three-phase power. Current transformers were placed in the discharge vehicle's explosion-proof box on the load side of the main breaker, thus coupling trans- ducers to line power and reducing current by a factor of 80. Voltage leads from the watt transducers were con- nected to the load side of the main breaker. Discharge Vehicle Power and Current Individual vehicle energy requirements are not neces- sarily one-twelfth of the total system power because of differences in instantaneous loading rate from one vehicle to another and also for the conditions that exists on a specific vehicle such as plugging or coal fines buildup on the conveyor deck. Therefore, the electrical power and current requirements of the discharge vehicle were singu- larly measured. A Rochester Instrument Systems watt transducer, identical to that measuring total system power, was used to measure the discharge vehicle power and a Transdata model 10CS501 current transducer was used to measure the current required by the discharge vehicle conveyor motor. Two 50:5 current transformers were placed in two of three conductors supplying power to the conveyor motor contactor, thereby coupling the trans- ducers to live current and reducing current by a factor of 10. Voltage leads were connected to the load side of the main breaker. System Voltage To monitor regulation of the power center and to en- sure that the system operating voltage remained within acceptable limits, the system voltage was measured using a potential transformer and a potential transducer. The 550:120 voltage tap was used on the Trenco TR12182 potential transformer to supply an acceptable voltage level to a Rochester Instrument Systems 10PS101. The trans- former primary was connected directly across the line on the load side of the main breaker. Output of this potential transducer was proportional to system voltage. Haulage Rate Perhaps the most important variable measured in the test program was haulage rate. It is the independent vari- able to which most other dependent variables are related. A Ramsey Engineering, model No. 10-20/40-20, belt weigh scale was used to measure instantaneous loading rate. Belt-scale electronics included an integrator to total the number of tons of coal that was conveyed. Conveyor-Reducer Drive Temperature Lubricant temperature of two conveyor-reducer drive gearboxes was measured to evaluate suitability of the gear- box for loads imposed on the system. Two thermocouple junctions were made from Type K thermocouple wire. Each was entered through a sealed fitting into the gearbox sump in the discharge and the 10th intermediate vehicles. An Omega model 2168 digital thermocouple readout was used to measure gearbox temperature. A Gould model 481 eight-channel strip-chart recorder was used to simultaneously record total system power, discharge vehicle power and current, system voltage, and haulage rate. Procedure Instrument Calibrations All sensors being recorded on the Gould strip-chart recorder were physically calibrated prior to the test program. In addition, prior to each day's run, the Gould recorder was calibrated with approximate voltages for each channel to simulate proper sensor stimuli using a General Resistance DAS66AX Dial-a-Source. Physical calibrations of the current, potential, and watt transducers were performed using a calibration device con- sisting of a bank of resistors (unity power factor) arranged to provide three-phase voltage and current to the trans- ducers. Voltage devices were connected directly across the line. Current was measured using a Fluke 80J-10 current shunt and Fluke 8600 DMM (digital multimeter). Voltage was also measured with the Fluke 8600 DMM. Transducer output was adjusted to the proper voltage level appropriate for input stimuli. Gain of the Gould recorder was adjusted to give appropriate scale deflection. These transducer voltages were inserted into the Gould recorder, via the Dial-a-Source, on a daily basis during the test program to maintain recorder calibration. The Ramsey belt scale was calibrated using weigh chains designed for that purpose. Calibration was per- formed with the belt running at normal operating speed with weigh chains placed on the moving belt. A three- point calibration was performed at 0-, 20-, and 40-pct full scale (0-, 4-, and 8-st/min loading rate). Gould recorder gain was also set to the appropriate level to indicate proper haulage rate. As with other transducers, the 23 recorder was calibrated daily using a simulated signal from the Dial-a-Source. Thermocouples were checked at ice point (32° F) using an ice bath and at boiling point (212° F) using a hotplate and water. Haulage Test A mixture of 2.5- by 2-in and 2- by 1.5-in coal from the HTRF was used in conveying tests from March 14 to April 5, 1984. For tests conducted April 5-12, 1984, ROM coal from the Bureau research mine was utilized. On March 14, 1984, first trial of the haulage test was performed. A Clark front-end loader was used to load coal into the HFB. Steady-state loading rate was adjusted from to 6 st/min over the test period from March 14 through April 12, 1984. Total system power, discharge vehicle conveyor motor current and power, system voltage, and loading rate were recorded on the Gould strip-chart recorder for all loading trials. On each day the testing was performed, the Gould recorder was allowed to warm up for 30 min and then it was calibrated with the Dial-a-Source. Belt-scale totalizer readings were recorded at the beginning and at the end of testing to determine the total amount of coal transported each day. A log was kept of all significant events, break- downs, and repairs during the test period. The log was voice-recorded on a microcassette tape recorder and later transcribed. Conveyor gearbox temperature was measured before the start of each test and every 10 min thereafter for the first 30 min. Subsequent readings were taken every 30 min or whenever the system was shut down. After completion of testing, coal carryback loss lying beneath each vehicle's conveyor drive shaft was weighed with a 1,000-lb capacity balance beam platform scale. Results and Discussion One objective of the MUCH system test plan was to demonstrate reliability of the conveying system by operat- ing for one full shift at 8 st/min, with no more than 24 min of downtime (95 pet availability). Because of a variety of recurring problems, no more than 136 min of running time was achieved in any one day (28 pet availability). Table 6 summarizes the amount of test time at the listed haulage rates. Table 6.-Haulage Av haulage rate, st/min (system emptv^ test summary Tesf time, min 197.0 1-2 244.2 3 65.0 4 23.9 5 29.0 6 1.5 Total 'seo.e ! 9 h 20.6 min. The maximum average conveying rate reached was 6 st/min. Conveyor tests were performed in 13 days from March 14 to April 12, 1984. Total test time was 560.6 min. Total run time achieved above 4 st/min was 30.5 min. During the test period, 682.1 st of coal was con- veyed. Average loading rate for the entire period was 1.22 st/min. Appendix B contains a conveyor test breakdown and repair log for this test period. The system test was started on March 14, 1984, at 2 st/min, but after 2 days, only 65 min of run time had been achieved because of breakdowns. The system was then allowed to run empty for 2 days to determine if failures were load dependent. After the system had cleaned itself out, the breakdowns ceased, but then it took a period of time for the coal to be cleaned out from beneath conveyor decks. It appeared that after rehandling the coal many times and by adding water with the water sprays, the coal became a fine, granular, dense mixture, which required considerably more energy to convey than dry, coarse coal. This wet mixture plugged up around conveyor sprockets, under the conveyor deck, and even on top of the conveyor deck. As coal was again added to the system, mechanical and electrical problems began to reoccur. Even as the loading rate was kept constant at 2 st/min, the rate of system malfunction increased with time. It is hypothesized that this was due to increased buildup of wet fines everywhere in the system. Shown in table 7 is a summary of results for these con- veyor tests. Discharge vehicle current, discharge vehicle power, and total system power are shown for loading rates from to 5 st/min. Figure 17 shows typical Gould strip- chart traces of real-time data. System power consumption ranged from 45 kW (60.3 hp) at no load (0 st/min) to 166 kW (222.5 hp) at 5 st/min. Average power consumption per vehicle 3.8 kW (5 hp) and 13.8 kW (18.4 hp) at and 5 st/min, respec- tively. Measurement of discharge vehicle power indicated that its power consumption ranged from 7.5 kW (10.1 hp) at no load to 10.0 kW (13.4 hp) at 2.2 st/min. This was greater than the discharge vehicle power measured at 5 st/min (8.9 kW or 11.9 hp), because wetted coal fines had built up on the conveyor deck to a depth of 4.5 in. Fig- ure 18 shows the extent of fines buildup on the conveyor deck. This buildup increased conveyor chain tension sig- nificantly and increased the coefficient of friction, requir- ing much greater power to convey the coal. Such buildup on the conveyor deck was also a problem with intermedi- ate vehicle 5. Once the coal fines buildup was removed from the deck, discharge vehicle power went down to 7.8 kW (10.5 hp), a reduction of 22 pet. Likewise, discharge motor current decreased from 21 to 14.8 A (25 pet decrease). Under no-load conditions, the system consumed up to 36 pet of the power required at 5 st/min. This was due mainly to the wet fines plugging up the conveying system, particularly material carried back under the conveyors in the vehicle hoppers. On March 16, 1984, after the system became plugged with coal and was emptied, residual 24 material caused the power consumption to remain artificially high at 68 kW (91.1 hp), even though no coal was being conveyed. On March 20, 1984, after allowing the system to run a period of time to purge itself, the power dropped 18 pet to 56 kW (75 hp). At that point, water sprays were turned on. After wetting the conveyor chains and deck for approximately 5 min, power consumption dropped another 18 pet to 46 kW (61.7 hp). One measure of conveying efficiency is the amount of energy required to convey material a given distance, expressed as specific energy in kilowatt hours per short ton. E *p = where E^ = specific energy, kW • h/st, P, = total system power, kW, and O = loading rate, st/min. O(60) In terms of energy requirements, most efficient conveying is achieved at the higher loading rates. At 4 st/min, specific energy was 0.41 kW • h/st conveyed. At the 1.8-st/min loading rate on March 15, 1984, the specific energy was 1.36 kW • h/st. Poorer efficiency was due to two factors: (1) material had built up on conveyor decks and underneath conveyors, and (2) there is a certain tare power requirement to overcome system friction under no- load conditions. Table 7. -MUCH haulage testing data summary-electrical parameters Haulage rate, st/min Discharqe convevor Current, Power, A kW System power, kW Total system voltage, V Coal conveying specific energy, 1 kW-h/(st-ft) Comments 3/14/84: 2.4 . . 13.5 15.5 13.0 17.7 17.0 14.0 19.8 15.2 NA NA NA NA NA NA NA 8.4 108 117 147 118 465 465 460 465 0.75 .85 1.36 1.23 Material had been on conveyor decks for several weeks prior to system startup. 2.3 . . 1.0 . . Values averaged over data 3.1 . . window. 3/15/84: 1.8 . . Material buildup on conveyor decks, mechanical problems occurred because of high loads. 15.5 st coal conveyed. 0.7 . . 4.0 . . 3/16/84: 1.6 3/20/84: . . . .. . 11.8 12.2 7.8 8.0 68 60 475 480 NAp NAp At 2: 11 p.m. At 3:11 p.m. 3/28/84: .. . ... . . . 12.0 10.2 12.2 7.8 7.5 7.8 56 46 45 485 485 485 NAp NAp NAp Before addition of water, 2:51 p m After addition of water, 2:56 p.m. At 3:14 p.m. 3/29/84: 2.0 13.0 7.6 87 475 .73 38.3 st coal conveyed. 3/30/84: 2.0 . . 11.5 13.0 21.0 7.4 7.4 10.0 81 99 138 475 470 460 .68 .41 1.05 4.0 . . 196.8 st coal conveyed. 4/05/84: 2.2 Discharge and intermediate vehi- cle 5 conveyor decks were plugged. 4/06/84: 3.0 14.8 10.5 138 460 .77 After-discharge and intermediate vehicle 5 conveyor decks cleaned. 59.8 st coal conveyed. 4/09/84: 5.0 18.0 8.9 166 455 .55 81.7 st coal conveyed. 4/12/84: NAp NAp NAp NAp NAp Mechanical failures and electrical problems effected the recorded data. 24.9 st coal conveyed. NA Not available. NAp Not applicable. ^otal mean-0.84 kWh/(st-ft); standard deviation, 0.30 kWh/(stft). NOTE.-ln tests conducted April 3 (44.9 st coal conveyed) and April 4 (46.6 st coal conveyed), electrical and mechanical problems were encountered. No operating data were obtained. 25 -to. LdCE gee ozO 0°- c?£z ok sco?, QO o £&> UJ x 150 100 50 20 10 250 12.5, March 14, 1984 Chart speed = 0.5 mm/s ■»%»i ■ to. .000 500 - 20 J 10 — I ■ 1 1 1- Ji'u-* • _ *~ 1 * ' _L i 1 1 1 . Discharge conveyor power no 100 operating 200 300 400 500 TIME, s 600 700 800 900 war . UZO QU ujrr.- £££< SOU | -R> to -I >-o to> UJ ^3. 300 200 100 March 30, 1984 Chart speed = 0.5 mm/s n 25 12.5 ,000 500 ■r^nr 1 ^T', vn '""Pf ' wwr-fy^-f^ f *t L Mull -i I i_ 20 \- E p 10 - ) 100 200 300 400 500 600 70 300 400 TIME, s 500 600 700 800 Figure 17.-Typical strip-chart data, 2.3 st/min (top) and 1 to 8 st/min (bottom). 26 Figure 1 8.-Coal fines buildup on conveyor deck. Throughout the test, the chain carried coal fines back underneath the conveyor decks. This material was ejected from the system at the point where the conveyor drive sprockets engaged the conveyor chain. It was only very fine material that piled up beneath the vehicles. It should be noted that no significant material loss was found around the hoppers at vehicle transfer points, even on those vehi- cles that were skewed at a steep angle. After this testing was completed, the amount of coal lost from each vehicle was weighed. Typical coal carry- back loss can be seen in figure 19. Table 8 shows the amount of coal lost under each vehicle and also quantifies that loss as a percentage of the total amount of coal con- veyed. As a percent of the total 682.1 st conveyed, loss from individual vehicles ranged from 0.026 pet in interme- diate vehicle 8 to 0.217 pet in intermediate vehicle 3. Average carryback loss was 0.063 pet per vehicle or 0.75 pet total. Vehicle 3 tended to lose more material than the others and was therefore manually cleaned out during the test. This in itself increased the amount of loss by allowing more room for material to drop from the convey- or. It should be noted that because material became finer as the test progressed, rate of carryback loss increased. In Table 8.-Conveyor carryback loss Vehicle Loss, lb Portion of total conveyed, 1 pet Lead 1 2 3 4 5 6 7 8 9 10 Discharge .... Total 'Total conveyed, 682.1 St. 614 554 747 2,957 755 426 475 514 355 885 1,241 740 0.045 .041 .055 .217 .055 .031 .035 .038 .026 .065 .091 .054 10,263 .75 an underground mine, coal will only pass through the system once. The percentage of fines will be much lower; therefore, carryback losses will be less. The conveyor gearbox temperatures of the discharge vehicle and intermediate vehicle 10 were monitored throughout the test program with Type K thermocouples. A temperature stabilization curve for gearbox lubricant is shown in figure 20. Data were obtained on March 28, 1984, at haulage rate of st/min for a 135-min period. A stabilization curve at various loading rates could not be obtained because the system would not operate for a sufficient period of time to stabilize oil temperature. The temperature at which the lubricant temperature stabilizes gives an indication of the load on the gearbox and its condition. Temperature in intermediate vehicle 10 stabilized at 131° F, with an ambient temperature of 56° F (or 77° F above ambient). This temperature is well within bounds of reasonable operating temperature. The discharge vehicle gearbox temperature did not stabilize during this period of time, but the slope of the curve was decreasing rapidly toward the end of the run. Temperature at the end of the run was 102° F above ambient. Under these conditions, discharge conveyor motor current and power consumption were 12 A and 7.8 kW (10.5 hp), respectively, at 0-st/min haulage rate. The results of the March 14 through April 12, 1984 conveyor tests follow: o A total of 682.1 st of coal was conveyed during 560.6 min of operation. o Coal fines, generated by recirculation of the coal, were getting wetter on each recirculation in the closed- loop system because of the water sprays. The wetted coal fines increased the operating loads significantly, which caused a high failure rate of conveyor drive components. Because of the large number of mechanical failures and electrical problems, no more than 136 min of operating time was achieved in any single day (28 pet availability). 27 o Conveying horsepower requirements ranged from 5 hp per vehicle at no load (0 st/min) to 18.4 hp per vehicle at 5 st/min (while conveying wet coal fines). o Specific energy required to transport the coal through the entire 228.8-ft system was 0.41 kW • h/st. o Almost no coal was lost through the system at vehi- cle transfer points. The only loss was through conveyor carryback, which represented about 0.75 pet of the total amount of coal conveyed. o Conveyor chain speed was 270.9 ft/min at 4-st/min haulage rate. Transport speed of discrete particles was 234 ft/min at 4 st/min. o Carryback on the conveyor return deck for each vehicle was between 0.5 and 1 in. Some material was consolidated and cemented to the deck, some was loose (fig. 21). o Very hard, consolidated material had built up in the pelican beak in front of the conveyor drive shaft (fig. 22). This area had been previously cleaned using a slate bar through the cleanout ports. This method of cleaning the pelican beak is ineffective, as the material becomes hard packed and will not flow out the cleanout ports. Even when loose, material will assume a natural angle of repose, building up to the edge of the cleanout ports before any coal exits the area. This material eventually is carried into the hopper. o The amount of coal in the hoppers varied from vehicle to vehicle. Hard-packed coal fines were found throughout the hopper, especially between the chain guides and sides of the hopper, of intermediate vehicles 3 and 4. It was evident that this probably created a bind in the system, increasing the operating loads. o There were not a great deal of consolidated coal fines in intermediate vehicle 8. o Coal fines were tightly cemented into place between the chain guides and sideplates on the return deck of some vehicles. Material was so hard that the chain rode over the top of the coal. o It was found in intermediate vehicle 3 that the chain guides were spaced too close together and the chain links, rather than the chain rollers, were riding on the chain guides. This also increased operating loads. Figure 19.-Typical coal carryback. 28 • Discharge vehicle conveyor gearbox A Intermediate vehicle conveyor gearbox 30 60 90 120 150 180 TIME FROM INITIAL STARTUP, min Figure 20.-Conveyor drive gearbox temperature rise. July 31 to August 3, 1984 Objective During previous haulage testing, it was found that one of the primary problems encountered was that of coal carryback. The coal carried back into the return deck increased power consumption and caused the system to periodically plug and shear drive pins, or cause other component failures. In an attempt to diminish the problems caused by carryback, it was decided to cut cleanout holes in the return deck to allow the coal carryback to exit, preventing its buildup. Therefore, cleanout ports were cut in the following areas of intermediate vehicles 3, 4, and 8 and the discharge vehicle: 1. Hopper- -Coal that is carried back into this area can cause the conveyor to stall, therefore two 4- by 8-in holes were flame cut just behind the foot shaft. 2. Pelican beak-Heavy buildup always occurs in this area. Even when the cleanout ports are removed, material will not clean itself from this area. Two 7.5- by 8-in cleanout ports were cut in the bottom of the pelican beak (fig. 23). 3. Return cfec/c.-Material builds up on this deck, in- creasing chain flight friction, therefore, the built-in outby cleanout port covers and middle cleanout ports were re- moved. The expectation was that the material that would otherwise be carried back would drop out onto the next outby vehicle hopper. 4. Area between chain guides and sideboard-Slots, approximately 0.75 by 6 in, were burned in the return deck between the chain guides and sideboards to allow material carried back by the chain links to exit. This area has a tendency to become plugged to a greater extent than the return deck. Coal buildup on the top decks was another problem that was addressed. Chain holddowns were fabricated from 2- by 2- by 0.25-in, 10-ft-long angle iron and bolted to the hopper sideboards of intermediate vehicles 3, 4, and 8, and the discharge vehicle. The chain guides on the return deck of intermediate vehicle 3 were removed, allowing the chain flights to con- tact the return deck. It was expected that this would help purge the return deck of coal. Procedure The MUCH system was trammed into a circle with the discharge vehicle dumping onto the lead vehicle, the bridge conveyor was not utilized. The belt scale was unavailable, as it was assigned to another project. Otherwise, instru- mentation that was used in April haulage test was again utilized. Total conveyor system power, discharge vehicle conveyor power, discharge vehicle conveyor current, and system voltage were measured on the Gould model 481 strip-chart recorder. The conveyor system was loaded with a fine mix of coal of an unknown size consistency from the Bureau's Hydraulic Transport Research Facility (HTRF) stockpile. Haulage rate was maintained at an estimated 3 to 6 st/min. Results and Discussion The system was operated for several hours during July 31 to August 3, 1984, at haulage rates between and 6 st/min (estimated). Because of the variability of the data and lack of haulage rate data, it was difficult to draw any definitive conclusions, but the system suffered no mechanical component failures (except shearpins) during this test period. 29 Figure 21 -Conveyor return coal buildup. # Figure 22.-Coal fines buildup in pelican beak. 30 Figure 23.-Cleanout ports In pelican beak. Generally, the steady-state power consumption was 85 to 95 kW, discharge vehicle current was 12 to 13 A, and power was 8 to 8.5 kW at an estimated (rough) 3 st/min. Comparing these figures to the regression equations devel- oped for discharge vehicle current (I d , = 16 A) and total system power (P„ 123 kW) during the April 1984 haulage testing, it appears that the cleanout ports were effective to some degree judging by the reduced motor current and power consumption. Upon teardown of the upper decks, it was observed that there was much less coal left on the return decks and in the hopper return area than in the previous haulage testing in March. The regression used were Discharge vehicle current (I d ) = 12.17 + 1.26 Q, and Total system power (P,) = 62.8 + 20.2 Q, where Q = haulage rate, st/min. The problem with these cleanouts was that an excessive amount of material was lost through the cleanout holes (fig. 24). Appendix C shows the carryback loss for each vehicle. Carryback loss averaged 264 lb for the unmodi- fied (no cleanout ports) vehicles. The modified (cleanout ports added) vehicles averaged a 2,255-lb loss, or an unacceptable 8.5 times greater carryback loss; therefore, the cleanout ports were covered. Chain holddowns were effective in preventing coal buildup on the top decks. None of the modified vehicles had any appreciable buildup on the top decks. The con- figuration of the leading edge of the holddowns on the discharge vehicle had to be modified. The ramped shape of the leading edge caused coal to wedge between the holddown and chain, pinching the chain, increasing sliding friction, and stalling the conveyor. Cutting off this ramped transition area resolved the problem, returning the motor current to its normal level. Removal of the chain guides on the return deck of intermediate vehicle 3 was very effective in cleaning the deck. 31 Figure 24.-Coal carryback loss after haulage testing, July 31 to August 3, 1 984. September 10-11, 1984 Objective Water sprays were fabricated to test the concept of spraying water in the return deck to minimize the problem of coal carryback plugging the system. Four spray bars or manifolds were fabricated (fig. 25) using eight Whirl jet 3/8 BD 8 size 1 nozzles manufactured by Spraying Systems Co. Two nozzles were used per spray bar, 16 in apart. Two spray bars were installed in intermediate vehicle 8 and two in the discharge vehicle. One spray manifold in each vehicle was inserted in the pelican beak through the cleanout ports. Nozzles were faced up and outby approxi- mately a 45° angle (fig. 26). The middle cleanout port in the return deck of both vehicles was removed and the manifolds were fastened into place with the nozzles facing up and outby. Cleanout ports used during the July 31 through August 3 haulage test were covered. Procedure The MUCH system was placed in a circle and instru- mented as before in July and August conveying tests. Testing was conducted as before, adding coal as needed to achieve an estimated haulage rate of between 3 and 6 st/min. Results and Discussion Water pressure at the machine inlet was 135 psi when the eight water sprays were functioning. Total flow rate measured was 2.69 gal/min or 0.336 gal/min per nozzle. The test did not run long before the coal was extremely wet, to the point of being a slurry. The material leaked out of the conveyors because of its liquid nature and the floor became extremely wet. This amount of water lost would present a serious problem underground, therefore the water spray manifolds were removed. Umbrella Miner Cutting Trials Support Objective To gain more information on the haulage characteristics of the MUCH system conveyors, it was decided to utilize the system to remove coalcrete cuttings that were pro- duced during umbrella miner cutting trials in the Bureau's Miner-Bolter Test Structure (MBTS). 32 Figure 25. -Water spray nozzles. Figure 26.-Water spray manifold in pelican beak. 33 Procedure The MUCH system was positioned to receive cuttings from the umbrella miner in the MBTS, to convey the material through the simulated workings, and then to discharge the cuttings from the bridge conveyor into the bucket of a front-end loader at the opposite end of the building. The system was instrumented to monitor the electrical system voltage, the total system electrical power in kilowatts, the discharge vehicle conveyor power in kilowatts, and the total system current requirements in amperes. Before testing, the conveyor system on each vehicle was run to verify proper operation and to remove any old coal fines from the conveyors. Each conveyor chain was oiled to free up any "frozen" links, and a new conveyor drive clutch was installed in the discharge vehi- cle. During this conveyor trial, coalcrete cuttings would be conveyed only once over the system, as opposed to the previous closed-loop type of haulage tests, which tended to overload the conveyors with a large quantity of fines from material degradation over a period of time. This trial would closely simulate actual coal haulage in an under- ground production situation. Results and Discussion Although the MUCH conveyor system performed al- most flawlessly during the trial and the only downtime was 10 min because of one conveyor drive shearpin failure, the trial did not produce much useful information. Mechanical difficulties with the umbrella miner while trying to cut the coalcrete were the main problems. An insufficient amount of coalcrete material was loaded to tax the MUCH con- veyor system much above its power requirements while running empty. The maximum power requirement observed during the trial, except for startup power, was 25 kW (33.5 hp) for the total system power. When the system was running empty with no load, 18.5 kW (25 hp) was required. The maximum observed loading was only a 35-pct increase over the no-load conditions. The maximum load required to start the system was 40 kW (54 hp). During the previous closed-loop conveyor trials con- ducted in 1984, no-load (0 st/min) power consumption data were taken for the system; these did not include the bridge conveyor. Because of previous coal haulage trials, a large amount of coal fines was in the return deck and pelican beak areas of the vehicles, as well as wet fines that had accumulated on the conveyor decks. No-load power consumption data during those trials ranged from 45 kW to 68 kW (60 to 90 hp), which is approximately three times higher than what was observed in this test. Obviously, the accumulation of coal fines and conveyor chain lubrication have a significant influence on the amount of power re- quired to drive the conveyor system and the subsequent loads. Conveyor System Problem Summary Upon reviewing the results of the four tests, it became obvious that there was a major problem with the conveyor system ability to handle fines. The coal fines, built up on the conveyor top deck and on the return deck, ultimately overloaded the system to the point of failure. The addi- tion of water only made the problem worse. Cleanout ports in the return deck helped reduce fines buildup but lead to excessive loss of material from the system by dumping fines on the bottom. The closed-loop test setup was also contributing to the fines handling problems. Conveying coal in a closed loop continuously degrades the coal and yields all fines, which is not reflective of a nor- mal ROM product that the system would be handling at an underground production face. Based on these obser- vations a number of changes were made to the system as follows. o A conveyor chain slack adjustment mechanism was added to each vehicle to provide a simple means of keep- ing the conveyor chains tight so as to limit fines buildup on the decks. This modification is described in the "System Modifications" section. o Autoguard torque clutches were removed from the conveyor drive systems and replaced by Lovejoy couplings. o The conveyor drive shearpins were removed and the shearpin couplings were welded solid. Conveyor drive motor thermal overloads were resized to adequately pro- tect the drive components. o The conveyor chain holddowns were removed, redesigned, and reinstalled. Approximately 12 st of 3- by 1-in gravel was purchased for further conveyor system evaluation after the comple- tion of these modifications. Approximately 5 st of the 3- by 1-in gravel was dumped onto the system. The system was started and ran a very short period of time before it shut down on electrical overload trippings because a conveyor jammed. The jam was caused when the conveyor chain rode over a rock (fig. 27) and the rock was caught between the tail sprocket shaft and the chain flight (fig. 28). The primary cause was the large gap between the end of the conveyor deck and the tail sprocket shaft as shown in figure 29. This gap permitted material to drop between the end of the conveyor deck and the tail sprocket shaft. Material that was too large to fall through would either be broken by the next flight and exit the gap, stay in the gap momen- tarily and lift the conveyor chain as it passed over (which allowed material to get caught under other chain flights), or it would become jammed between the chain flight and the tail sprocket shaft. Fines that were small enough to 34 fall through the gap would mostly fall into the hopper of the next vehicle but a fraction would be carried by the return chain flights into the return deck to cause further jamming problems. The solution to these jamming problems was to (1) de- sign and install a plate on the conveyor deck to bridge the gap, and (2) redesign the conveyor chain holddowns on each vehicle. These modifications are fully described in the "System Modifications" section of this report and are shown in figures 30 and 31. Coal Conveyor Acceptance Test-October 29, 1986 Upon the completion of the conveyor modifications on all vehicles, except the bridge conveyor, the system was run empty for a couple of hours to check operation and to break the system in before attempting to load the conveyors. After the initial break-in of the conveyors, the system was moved outside to a level open area and was trammed into a closed loop for testing. The bridge con- veyor was included in the loop. Approximately 8 st of 1- by 3-in gravel was loaded onto the conveyors and the conveyors were run in a closed loop for 30 min with no problems occurring. The rock was then conveyed off the system and 6 st of ROM coal was loaded onto the system for the next day's acceptance test. Objective The goal of the conveyor acceptance test was to con- tinuously convey coal at a rate of 8 st/min over an 8-h period while maintaining a system availability of 95 pet. For an 8-h test (480 min) this allowed only 24 min of downtime. Figure 27.-Conveyor chain riding over rock. 35 Figure 28.-Rock caught between shaft and flight. Figure 29.-Gap between deck and shaft. 36 Figure 30.-Conveyor deck plate. Figure 31 .-Conveyor chain holddowns. 37 Procedure DRAWBAR PULL TEST The test was conducted on October 12, 1986. The log of the test event is given in appendix D. The test was initiated with 6 st of ROM coal on the system. During the first 30 m in of the test an additional 6 st of coal was added to make up for the loss of fines as the coal degraded and to maintain 8 to 10 st of material on the system. A belt scale was not utilized during the test to avoid unnecessary downtime the additional belt might cause. Approximately 1 h into the test an additional 2 st of coal was added to the system. At this point in the test a total of 14 st of ROM coal had been utilized, approximate- ly 12 st was on the system and about 2 st had been lost as fines. As the test continued the coal kept degrading and fines were being lost, so after about 2 h of running time, 3 st of the 1- by 3-in gravel was added and mixed with the coal. Because the test was being conducted outside, and to avoid problems previously encountered because of excess water, very little water was added to the material during the test. The water sprays were used only when the dust became excessive. As the test continued, the conveyor drive motors thermal overloads kicked out because of excessive loads, especially on the lead vehicle. The haulage rate was estimated between 12 to 13 st/min of mixed coal and rock. Approximately 2 st of material was removed from the system and the test continued. The test was terminated at 2:24 p.m. with over 8 st of material remaining on the system. Results and Discussion The test time was from 8:00 a.m. to 2:45 p.m., a total of 405 min. The total haulage time was 337 min and the total downtime was 68 min. The system availability for the test was 83.2 pet. The majority of the downtime, 58 min or 85 pet, was due to thermal overload trips on the con- veyor drive motors on the lead, intermediate vehicle 2, and intermediate vehicle 5. After the test the conveyor drive motor thermal overload setting on each vehicle was checked. On the three problem vehicles the settings were found to be lower than on the other vehicles, which con- tributed to the excessive tripping problem. The total amount of coal-rock mixture conveyed is estimated by 337 min of haulage time at an approximate rate of 10 st/min or 3,370 st. A total of 17 st of material was loaded on the system of which 2 st was removed and 8 st remained on the system at the end of the test. There- fore, 7 st of material was lost during testing, mostly as fines, with some of the loss due to spillage at dumping points between vehicles. The total amount of lost mate- rial, 7 st, is equivalent to 0.2 pet of the total 3,370 st conveyed. Objective A drawbar pull test was conducted to measure the trac- tive effort produced by the MUCH system. The amount of tractive effort available gives an indication of the capability of the system to tram up grades and, if needed, to tow or pull a disabled or buried piece of equipment. It should be noted that the MUCH system is not intended or designed to be a towing vehicle and the use as such could lead to system damage. Procedure The drawbar pull of the MUCH system (without bridge conveyor) was measured by pulling against a 50-st-capacity Dillon dynamometer, which was anchored to a 35-st mobile crane (fig. 32). The dynamometer was attached to the discharge vehicle at the conveyor bridge mounting holes in the frame using a heavy chain. The dynamometer has a resolution of ±500 lb and a factory-stated accuracy of 0.5 pet of full scale. The brakes of the crane were applied during the pull test to keep the crane from mov- ing. The test was conducted outdoors on a heavily com- pacted, level, dry, dirt surface. The system was trammed and positioned into as straight a line as possible prior to being connected by chains to the dynamometer and crane. Results and Discussion The first pull test was conducted with no material in the conveyors. The empty weight of the MUCH system is approximately 117,800 lb. The tractive effort measured by the dynamometer was between 27,000 and 30,000 lb of steady pull. It was observed during this trial that some of the tires were spinning against the ground and on some of the intermediate vehicles the wheel rims were spinning inside the tires. This may be one disadvantage of having foam-filled tires as opposed to air-filled tires, the adhesion of the foam-filled tire to the rim is less in some cases than the friction between the tire and the ground, which can reduce tractive effort. The second pull test (fig. 33) was conducted after ap- proximately 10 st of material was loaded onto the MUCH system, which increased the total weight to approximately 137,800 lb. The steady drawbar pull increased to about 35,000 lb. After a few moments of tractive effort at this level three of the tram motor overloads tripped, shutting tram motors down. 38 Figure 32.-Crane and dynamometer used during drawbar pull test ROUGH TERRAIN TEST Objective Figure 33.-Drawbar pull test A rough terrain test was conducted to evaluate the performance of the system while tramming over areas with difficult bottom conditions. Procedure The test was conducted outdoors in an area with a surface of mixed mud and coal that had been saturated with water. A 2-ft-high by 13-ft-wide by 13-ft-long pile of mixed dirt, coal, rock, and 4-in by 4-in by 4-ft wood crib blocks was constructed as an obstacle. A roughly 10-ft- diam, 6-in deep water-mud hole was also part of the test course. The MUCH system was trammed repeatedly over the test course. Results and Discussion The system trammed over and through the pile of material and water-mud hole with no problems and no vehicle-to-vehicle binding or interference was observed. The front hopper on the lead vehicle acted as a plow to 39 help remove high obstacles prior to driving over them with the system. The wet, muddy, slippery bottom conditions had no adverse effect on the tracking ability of the system. As tramming over the test course continued, 6- to 8-in- deep ruts developed in one area that were of such length that two vehicles had their drawbars and undercarriages dragging through the mud. There was no noticeable loss of tram speed or tracking ability in this area. CANOPY LOAD TEST The operator canopy, which protects the lead vehicle operator, was redesigned and load tested. The canopy redesign and installation is described in the "System Mod- ifications" section of this report. As stated in the Code of Federal Regulations (30 CFR 75.1710-1), a cab or canopy must have a "minimum structural capacity to support elastically: (1) A dead weight load of 18,000 pounds, or (2) 15 p.s.i. distributed uniformly over the plan view area of the structure, whichever is lesser." The elastic load criterion required by Federal law is based upon a combined statistical, analytical, and experi- mental investigation where it was found that a cab or canopy designed to meet this elastic load criterion has at least enough potential energy (in the form of available strain energy) to withstand the majority of roof falls as determined by the statistical analysis of all fatal roof falls from 1966 to 1972. The new MUCH system canopy was tested following MSHA guidelines. 7 The following items were used during the canopy load test: 1. A Sensotec, Inc., 20,000-lb-capacity compression force transducer. 2. A Fluke digital multimeter. 3. A Honeywell Accudata 218 bridge amplifier. 4. A 20-st-capacity hydraulic jack. 5. A displacement dial indicator accurate to 0.001 in. 6. A 12-in I-beam, 24 in long. 7. The roof cart of the miner-bolter test structure (MBTS). 8. Wood cribbing. Sawyer, S. G., and D. K. Brogam. A Testing Procedure for the Certification of Underground Protective Cabs and Canopies. MSHA IR 1002, 1974, 15 pp. Procedure In general, the test procedures, as outlined in MSHA IR 1002, were followed. Specifically, the lead vehicle on which the canopy is mounted was positioned directly under the roof cart of the MBTS and wood cribbing was used to support the canopy base and raise the lead vehicle's wheels off the floor. The top of the canopy was marked to show the middle ninth area of the total plan view area and the centroid of the canopy top was marked on the bottom sur- face of the canopy. A 12-in I-beam was cut and machined to cover and distribute the load over the middle ninth area of the canopy. The Sensotec 20,000-lb-capacity compres- sion force transducer, which was previously calibrated, was placed on the I-beam directly above the centroid of the canopy and connected to the Fluke digital multimeter and the bridge amplifier. The 20-st-capacity hydraulic jack was placed on the force transducer and also made to contact the bottom of the MBTS roof cart. The displacement dial indicator was mounted on the floor of the operators sta- tion and in contact with the centroid location on the underside of the canopy. The testing was conducted by utilizing the hydraulic jack to load the canopy in 1,000-lb increments, by reacting against the MBTS roof cart and by using the dial indicator to measure the canopy deflection. Once the canopy was loaded to 18,000 lb the load was removed and the residual deformation of the canopy was measured. Results and Discussion The results of the load test are shown graphically in figure 34. The maximum canopy deflection at a load of 18,000 lb was 0.152 in at the centroid and the residual deformation at the centroid, once the load was removed, was 0.009 in. This residual deformation is 5.9 pet of the total deflection and falls well within the maximum allow- able residual deformation of 10 pet as stated in MSHA IR 1002. Therefore, the canopy is classified as substantial and is certifiable by a site-registered engineer. GROUND PRESSURE EVALUATION Objective A ground pressure evaluation was conducted to deter- mine the amount of pressure (pounds per square inch) that will be applied to the mine floor by the MUCH system tires during normal underground operations. This value is important in that the higher the pressure is at the tire-floor interface, the more likely the floor will be to deteriorate and ruts will be generated. Results For the evaluation, an average tire "footprint" was determined to be 10 in long by 8.5 in wide or 85 in 2 . Each vehicle has four tires for a total ground contact area of 40 340 in 2 . The unloaded weight of the lead vehicle was 11,600 lb, the discharge vehicle weight was 13,200 lb, the intermediate vehicles weighed 9,300 lb each, and the bridge conveyor weight was about 3,000 lb. The ground pressure per tire for each vehicle empty and loaded with 1 st of coal per vehicle is given in table 9. The ground pressure is the highest at the discharge vehicle with the bridge conveyor attached. Table 9.-Tire ground pressure for loaded and unloaded vehicles Vehicle Weight, lb Rear tire ground pressure, psi Lead: Unloaded 11,600 34 Loaded 13,600 40 Intermediate: Unloaded 9,300 27 Loaded 11,300 33 Discharge: Unloaded 13,200 39 Loaded 15,200 45 Discharge and bridge: 1 Unloaded 14,700 48 Loaded 17,700 54 *Front tire ground pressures were 39 psi (unloaded) and 45 psi (loaded). SURFACE TEST SUMMARY The MUCH system was evaluated and modified during the surface test program to prepare the system for an in- mine production trial. The following are highlights of the surface test program. o The 12-unit MUCH system successfully trammed inby and outby through a simulated underground work area with 20-ft-wide entries and crosscuts at 90° to each other and 48-ft-square pillars. During these tracking and retracking trials, the system tended to drift toward the inside radius of the 90° turns. When tramming in the inby direction the maximum observed variance of any vehicle from the path of the lead vehicle was 68 in. When tramming in the outby direction the maximum observed variance of any vehicle from the path of the discharge vehicle was 101 in. The system tracked and retracked with enough consistency to operate in an underground mine of similar dimensions. o A maximum tracking variance of 35 in was observed during the simulated production cycle trial when the system was operating behind a simulated continuous miner making a 20-ft cut, a lift change, and second 20-ft cut. o The time to complete a simulated place change was 3 min 20 s while tramming over a distance of 245 ft. KEY Load, Displacement, I0 3 lb in 0.000 1 .007 2 .017 3 .027 4 .036 5 .044 6 .053 7 .062 8 .072 9 .076 10 .086 1 1 .095 12 .103 13 .120 14 .123 15 .127 16 .136 17 .142 18 .152 ■ Max deflection = 0.152 in • Residual deflection = 0.009 in 0.04 0.08 0.12 0.16 DEFLECTION, in Figure 34.-Lead vehicle operator canopy load test, load versus deflection. o A noise level survey was conducted. At the lead vehicle operator compartment, a noise level of 102 dBA was measured when the conveyor system was operating with no coal. At a haulage rate of 3.7 st/min of coal, 98 dBA was measured; at a haulage rate of 7.4 st/min, 97 dBA was measured at the operator station. The noise levels measured indicate that some form of hearing pro- tection will be required for the MUCH system operators to maintain their 8-h noise exposure to 90 dBA or below. o The conveyor chain speed was measured on a num- ber of vehicles. The average conveyor chain speed with no load was 273 ft/min. The chain speed at a haulage rate of 7 st/min was 271 ft/min. The average startup time for the conveyor system was 1.36 s at no load and 1.86 s at a 4-st/min load rate. o After a number of conveyor tests and numerous modifications on conveyor system, an acceptance test was conducted. During the test, the conveyor system operated for 337 min out of a total test time of 405 min (83.2 pet availability), the haulage rate averaged 10 st/min of mixed coal and rock, and the total amount of material conveyed was approximately 3,370 st. 41 o A drawbar pull test was conducted in which the empty MUCH system exerted a steady pull of 30,000 lb tractive effort; when the system was loaded with 10 st of material, the drawbar pull increased to 35,000 lb. o During a rough terrain test the system trammed through a 2-ft-high mixture of mud, coal, rock, and crib blocks, a 6-in-deep water-mud hole, and numerous 6- to 8- in-deep ruts with no loss in tram speed or tracking ability. o The redesigned operator canopy was load tested with a static load of 18,000 lb applied to the center ninth area of the canopy. The maximum deflection was 0.152 in and the residual deformation once the load was removed was 0.009 in. The canopy is classified as substantial and is certifiable by a State registered engineer. o A ground pressure evaluation was conducted. The highest ground pressure was at the rear axle of the dis- charge vehicle with the bridge conveyor attached and a full load of coal on the discharge vehicle and bridge con- veyor. The maximum rear tire ground pressure was 54 psi. SYSTEM MODIFICATIONS Numerous modifications were made to the MUCH sys- tem to prepare it for in-mine trials and to correct observed deficiencies found during surface testing. A summary of the modifications is given in appendix F. The system modifications are discussed in the following sections. MSHA EXPERIMENTAL PERMIT APPROVAL In order to allow an in-mine trial of the MUCH system, an experimental permit had to be acquired from the Mine Safety and Health Administration (MSHA) Approval and Certification Center. The effort to acquire an experiment- al permit was initiated by a meeting of Bureau, Boeing Services, and MSHA personnel on January 24, 1985, at the MSHA Approval and Certification Center, Triadelphia, WV. The purpose of the meeting was to discuss the pro- cedures of permit acquisition, to find out what materials MSHA currently had concerning the MUCH system (JMMD had at one time initiated an approval effort), and to acquire information on any new changes in permit requirements that might effect the system. The formal application for the experimental permit was sent to MSHA under Company Application Code No. 112525 on July 29, 1985. After numerous drawing changes, many MUCH system revisions, and much correspondence with MSHA, experimental permit approval was received for the system on October 1, 1986. Experimental permit No. EP-541-0 as issued for the MUCH system had a duration of 6 months and with a request to MSHA, was renewed on April 7, 1987, for an additional 6-month period. During the approval process, the following electrical deficiencies were defined and resolved to meet MSHA's requirements. The electrical control circuits of the MUCH system, as designed by JMMD, operated at a voltage of 460 V ac line-to-line. As stated in a MSHA memorandum, dated October 25, 1982, "The voltage of alternating current control circuits shall not exceed nominal 120 V line-to- line." The 460-V ac MUCH control circuits were in violation of MSHA regulations and had to be dropped to 120 V to be approvable. The electrical control circuit was modified by installing a Mining Control, Inc., 24610 460-V primary to 120- V secondary, 2-kV«A, stepdown control transformer contained in an approved (X/P 2504-1) enclo- sure mounted on the discharge vehicle adjacent to the system's main electrical control box. The low-voltage secondary was equipped with 15-A fuses to protect the control circuitry wiring. The control voltage reduction necessitated replacement of the main circuit breaker (CB- 1), shunt trip coil, and all magnetic coils for the motor control contactors and time delay relays. The emergency stop circuit was not-fail safe because normally open contacts at the emergency stop switches had to be closed to activate the shunt trip coil and open the contacts of the main circuit breaker. If the emergency switch contacts were not functional because of corrosion or mechanical problems, the circuit could not be completed to activate the shunt coil and open the breaker contacts. The emergency stop circuit and shunt trip coil were changed so that a closed circuit was required to activate the shunt trip coil and hold the main circuit breaker contacts closed. The contacts of emergency stop switches were changed from normally open to normally closed so that when the switches were activated, the contacts would open and break the circuit to the shunt trip coil and open the main breaker contacts. The trailing cable of the MUCH system was originally a 1/0, three-conductor, 90° C cable rated at an approxi- mate ampacity of 219 A for underground use. The total full load motor current of the system is approximately 374 A and a normal load that would be expected during underground operations over a given period of time would be approximately 250 A. Obviously, the 1/0 trailing cable was of insufficient ampacity to meet MSHA requirements. The 1/0 cable was replaced with a three-conductor, 3/0 G- GC, 2,000- V, 90° C cable with a maximum ampacity of 294 A. The 1/0 size power cable between the discharge vehicle and intermediate vehicle 10 was replaced with 3/0 cable; and 1/0 power cables between intermediate vehicles 10 and 9 and between 9 and 8 were replaced with 2/0 cable. 42 The electrical fault ground-check circuit was original- ly connected to grounding studs in the explosion-proof (XP) box of each vehicle. This was not an MSHA ap- proved circuit because an open ground fault circuit to an individual vehicle would not be detected if the ground- check circuit to any other vehicle was continuous. The circuit was revised such that the ground check could only be connected to the ground stud in the lead vehicle, thereby providing one continuous path instead of a parallel path at each vehicle. The headlight circuit on the discharge vehicle was originally energized through a connection box located on the right side of the vehicle, which did not meet MSHA specifications. The circuit was changed to include a two- pole lever action pushbutton switch mounted in an XP enclosure in lieu of the connection box. Inspection of the MUCH electrical system components revealed that all electrical motor thermal overload relay heater elements were improperly sized to adequately protect the motors during normal operation. All heater elements were replaced with properly specified elements. The following refurbishment, replacement, and/or repair actions were required prior to the MSHA on-site inspection. 1. Electrical cable conduit and clamps were installed or relocated to provide sufficient protection to electrical cables. 2. The routing and suspension of electrical cables between vehicles was modified to prevent mechanical damage. 3. The trailing cable ground-check circuit was checked for proper operation. 4. Certification tags on electrical components were replaced or relocated to permit easier access during field inspection. 5. The cable for the pager phone system was inspected and repaired as necessary. 6. All hose conduit damaged during the test program was replaced or repaired. 7. All cable packing glands were checked for proper clearances. 8. All control box covers and edges were polished and checked for proper clearances. 9. All electrical components in control boxes were tagged for easy identification. 10. All electrical connections and wire terminals were checked and tightened as necessary. 11. Electrical layout and wiring diagrams were updated to include all electrical modifications and corrections. OPERATOR CANOPY REDESIGN The operator canopy, which protects the lead vehicle operator, was redesigned and load tested. During tram- ming operation of the system in the EMTA, the top of the canopy came into contact with the simulated ribs of the EMTA thereby bending the canopy's vertical supports (fig. 35). Although this original JMMD canopy design met MSHA requirements for vertical load capacity, it was very weak under horizontal loading conditions. After the three vertical supports were deformed by rib contact, the ability of the canopy to support the MSHA-required vertical load was reduced to an unacceptable level. This required rede- sign of the canopy to increase the strength and number of canopy supports. In the new design (figs. 36-37), the three vertical sup- ports were replaced by four vertical supports. The new supports are constructed of rectangular steel tubing instead of threaded rod as was used in the original design. Each of the four vertical supports consists of a length of 4-in- long by 2-in-wide by 1/4-in-thick rectangular steel tubing welded to the canopy and a length of 5-in-long by 3-in- wide by 3/8-in-thick rectangular steel tubing welded to the base of the operator compartment. The 4- by 2-in tubing fits inside the 5- by 3-in tubing and is secured in place by locking pins inserted through holes drilled in both lengths of tubing. The modified canopy has an operating height range of 42 in to 56 in, and is adjustable in 3.5-in incre- ments over the height range. The canopy top strength was increased by using 1/2-in-thick steel plate instead of the original 3/8-in plate and by increasing the size of the steel tubing frame. Additional structural modifications were made to increase the strength of the floor area. Upon the completion of the installation of the modified canopy the structure was load tested and met MSHA requirements. ELEVATION OF CONVEYOR DECK In the original JMMD design there was approximately 1.5 in of clearance between the forward edge of a vehicle's coal receiving hopper and the bottom surface of the con- veyor deck of the next forward vehicle when the system was on level terrain, as shown in figures 38 (top) and 39 (top). During tramming and maneuvering of the system it became obvious that additional clearance was needed in this area. When tramming over uneven terrain, especially when one vehicle is twisting along the longitudinal center line relative to the next as when one wheel runs over a crib block, the vehicles bind together in this throat area (fig. 40, top) and the tracking ability of the vehicle is lost. To alleviate this problem, the frame of each vehicle was modified to raise the discharge end of the conveyor deck by approximately 5 in. This modification increased the 43 height of the throat area from 13 in (fig. 40, top) to 18 in (fig. 40, bottom) and increased the hopper clearance from 1.5 in to 8 in (figs. 38-39, bottom). CABLE HANDLING TRAY Cable handling trays were fabricated and installed through the drawbars between vehicles. The trays (fig. 41), fabricated from 8-in C-channel, pivot in the middle to help reduce cable and hose damage. The trays support the power cable, control cable, telephone cable, and water hose between vehicles. DISCHARGE VEHICLE STEERING SYSTEM The hydraulic steering unit on the discharge vehicle was inadequate to allow the system to tram through the EMTA in the outby direction. A greater steering angle was needed; therefore, the linkage system was redesigned and modified to increase the steering angle and still utilize the existing cylinder. A bell crank arrangement was manufac- tured, and the steering cylinder was relocated. Figure 42 shows the arrangement. EMERGENCY SHUTDOWN SYSTEM During the surface tests it became obvious that an improved method of emergency shutdown of the MUCH electrical system was required. The system as received had only three locations of emergency shutoff, one at the lead vehicle operator station and one on each side of the dis- charge vehicle; these two locations are approximately 225 ft apart. In an underground situation neither the lead nor discharge vehicle operator can see the middle third of the system. If a problem occurred at a location not visible to either operator there was no way to shut down the entire system. To remedy this situation during surface testing, an emergency shutdown pull cord was installed along the center of the system for the entire length, which allowed complete system shutdown from any location (fig. 43). It must be noted that although this shutdown system worked well during the surface testing of the MUCH system, it was not designed for underground use. Figure 35.-Original operator canopy. 44 Figure 36.-Lead vehicle operator canopy. Figure 37. -Operator in lead vehicle. 45 «^»* Figure 38.-Clearance between hopper and deck-original Figure 39-Conveyor deck. Before modification, 1 .5-in throat design (top) and modified design (bottom). clearance (top); after modification, 8-in throat clearance (bottom). r I \r\rtr A/ Figure 40.-Height of throat area-original design (top) and modified design (bottom). 46 Figure 41 -Cable handling tray. Figure 42.-Discharge vehicle hydraulic steering system arrangement 47 Figure 43.-Emergency shutdown pull cord. CONVEYOR CHAIN SLACK ADJUSTMENT MECHANISM During coal haulage trials, problems occurred from fines buildup on the conveyor decks, which significantly increased haulage power requirements. One major cause of these problems was the slack in conveyor chain. As the MUCH system was originally designed, the only way of removing slack from the conveyor chain was by physically removing one or more links from the chain; but this could only be accomplished when 2 in or more of slack was available. Under extreme circumstances 2 in of chain slack was sufficient to allow a buildup of approximately 5 in of fines on the conveyor deck. To correct this problem, a conveyor-chain slack adjustment mechanism was designed and installed on the vehicles. The mechanism, as shown in figures 44 and 45, consists of two adjustment screws-one on each of the conveyor sideboards. The sideboards, which support the tailshaft of the conveyor chain, were cut approximately 65 in from the tailshaft, and the sideboard mounting holes were slotted to allow 5 in of movement. The adjusting screws are mounted on the forward half of the sideboard, and used to position the rear half of the sideboards and tailshaft to adjust conveyor-chain tension. The sideboard mounting bolts hold the sideboards and tailshaft in position when the chain adjustment is completed. CONVEYOR CHAIN HOLDDOWNS In the course of evaluating the conveyor system it be- came obvious that an effective conveyor chain holddown system was needed to help prevent fines buildup on the conveyor decks. Prior to the final conveyor acceptance test, a chain holddown system was installed on all vehicles (fig. 31). The holddowns consisted of 2- by 2- by 0.25-in steel angles bolted to the conveyor sideboards with 0.25- in clearance between the top of the chain and the bottom of the angle. The holddowns extend the entire length of the conveyor deck from the end of the loading hopper to about 4 in past the tail sprocket, which minimizes the amount of material that could be pulled in between the chain and the holddowns. A 2.5- by 0.25-in steel strip was welded between the legs of the angle to eliminate another potential area of material buildup. The chain holddown mounting holes in the rear portion of the con- veyor sideboard were slotted to allow chain slack adjust- ment without removing the holddowns (fig. 46). 48 Figure 44. -Conveyor chain slack adjustment screw. IS O O QifcuufiQipi ifly it iu)n m ( l"l C »»■■' '"■■■ ■ " ■■£ ^Q "'Hill, 111 IN PLAN r-^5-^ Center line -"fl 1 - QEHOfiggl iJii'm-iji'i' i'B -w* igr- "#" ELEVATION Figure 45.-Conveyor chain slack adjustment mechanism. 49 Figure 46.-Chain holddown mounting hole. Figure 47.-Conveyor deck extension plate. 50 CONVEYOR DECK EXTENSION As discussed in the "Conveyor System Tests" section, there was a material conveying problem because of a gap between the end of the conveyor deck and the tail sprocket shafts (fig. 29). This gap allowed chunks of material to get caught and either jam the conveyor or elevate the chain. This problem was amplified by the addition of the chain tensioning system by which the chain tension was increased by moving the sideboards. Increasing the chain tension also moved the tail sprocket shaft away from the conveyor deck, increasing the size of the gap. This problem was corrected by the addition of a 0.25-in-thick steel plate to bridge the gap between the conveyor deck and the tail sprocket shaft (fig. 47). The plate lies on the conveyor deck with the rear edge as close as possible to the tail sprocket shaft. The two outside edges of the plate are welded to the two conveyor sideboards so that the posi- tioning between the plate and the sprocket shaft remains constant when the conveyor chain tension is adjusted. The leading edges of the plate are chamfered so as not to bind the conveyor chain. CONCLUSIONS Surface testing showed that the MUCH system has the potential to substantially increase the productivity of a room-and-pillar mining system. The following items are recommended to improve its general functioning in an underground operation. o An improved emergency shutdown system should be designed and installed on the system to allow electrical deactivation from any point along either side of the con- veyor train. o Voice-activated headsets should be used by the lead and discharge vehicles operators to permit hands-off communications. o State-of-the-art noise reduction technologies should be applied to reduce the conveyor system noise levels. o MSHA-approved in-line cable connectors should be installed in the power and control cables between vehicles. These connectors would reduce vehicle change out time. o All conveyor drive components should be strength- ened by at least 50 pet. o A positive means of locking the foam-filled tires to the rims is needed to prevent the tires from spinning on the rims. o A cable reel mounted along the panel belt for the MUCH system trailing cable would help to limit the possi- bility of the discharge vehicle tramming over the cable. o An adequate supply of spare parts should be avail- able at the mine site to minimize downtime. Appendix E is a recommended spare parts list. o A thorough operator's training program should be conducted for the face production personnel prior to the start of an in-mine trial. o Additional lighting along the length of the system would be an advantage for both tramming and safety. o Audio signal on the system to warn the persons when it moves forward or rearward. o Rearview mirror in the lead vehicle operator cab to permit the operator to see behind the lead vehicle without turning around. 51 APPENDIX A.-MUCH SYSTEM SPECIFICATIONS The MUCH system is designed to be used in an underground room-and-pillar mining system. Because it is a versatile system, it can be used in a highwall operation also. Total system length (12 units) ft . . 250 Estimated total weight (12 units including bridge conveyor) lb . . 120,800 Power, 3-phase, 60 Hz V ac . . 460 Minimum turning radius ft . . 24 Mine configuration: Entry-crosscut width ft . . 18-20 Entry-crosscut angle deg . . 60-90 Minimum working height, in: With canopy on lead car 60 Without canopy on lead car 48 Vehicle units Lead vehicle Intermediate vehicle Discharge vehicle Frame: Length ft Active length ft Height in Width ft Canopy (adjustable) in Conveyor: Chain speed ft/min Chain width in Trough height in Motor hp Capacity st/min Tram: Speed ft/min Motor hp Wheel size Tread width ft Brakes Drive Steering: Forward Reverse Communications Hydraulics: Pump Motor hp Headlights: Number Voltage V ac NAp Not applicable. '28.25 ft with bridge conveyor. *22 ft with bridge conveyor. 23.25 21.75 '21.75 19.75 19 2 19 41 44 41 6.5 6 6 42-56 NAp NAp 280 280 280 30 30 30 9 9 9 15 15 15 12 12 12 80 80 80 7.5 5 7.5 8.25 by 15 8.25 by 15 8.25 by 15 5 5 5 Disk, hydraulic Spring activated, electrically Disk, hydraulic released Front wheel Front wheel Rear wheel Manual Automatic NAp Automatic NAp Manual Pager phone Optional on 1 Pager phone unit 1 unit NAp 1 unit 1 NAp 1 2 None 2 11 NAp 11 52 APPENDIX B.-CONVEYOR TEST BREAKDOWN AND REPAIR LOG Cumulative run time, Av loading rate, Machine mm st/min 3/14/84: 10 ... 2 MUCH 18 ... 2 HFB 50 ... 2 HFB 3/15/84: 59 ... 2 MUCH 59.1 . 2 2 MUCH 63.1 . MUCH/HFB 64.1 . 1 MUCH 65.1 . 1 MUCH Description 3/16/84: 68.6 1.5 1.5 MUCH 69 MUCH 76 1.6 MUCH 3/20/84: 79 MUCH 3/28/84: 215 3/29/84: 283 2 MUCH 284 2 MUCH Breaker on intermediate vehicle 8 tripped out, reason unknown. Breaker was reset. Main breaker on HFB tripped out. Unable to reset breaker. Problem was diagnosed as a poor connection on source side of the breaker. Connections were tightened, resolving the problem. Main breaker on HFB tripped out, reason unknown Breaker was reset. Clutch slipped on intermediate vehicle 10. Vehicle was jogged back and forth to reset the clutch and unjam the conveyor. Intermediate vehicle 9 breaker tripped, reason unknown. Breaker was reset. Power center tripped out, reason unknown. Breaker was reset. Intermediate vehicles 2 through 10 and the discharge vehicle breakers tripped out, reason unknown. Intermediate vehicle 9 sheared a pin. Shearpin was replaced and the breakers were reset. Key sheared on conveyor drive shaft on intermediate vehicle 9. Keyseat was damaged and driveplate bore was galled. To repair the conveyor driveshaft, the driveplate was welded solidly to the driveshaft. Clutch slipped on intermediate vehicle 10 and was reset. Clutch slipped on intermediate vehicle 10 and was reset. Breakers on intermediate vehicles 7 and 8 tripped out, reason unknown. Intermediate vehicle 9 sheared a pin. Breakers were reset and shearpin was replaced. Discharge vehicle speed sensor shaft seized and failed the stub shaft on end of tailshaft (speed sensor sprocket drive). Speed sensor was electrically bypassed. Replacement speed sensor and tailshaft were later replaced. No breakdowns. Breaker on intermediate vehicle 4 tripped out, reason unknown. Breaker was reset. Breaker on intermediate vehicle 8 tripped out, reason unknown. Breaker was reset. See explanatory notes at end of tabulation. 53 Cumulative run time, min Av loading rate, st/min Machine Description 3/29/84: 298 .. 3/30/84: 299 .. 341 .. 341.5 . 346.5 . 360.5 . 362.5 . 4/03/84: 364 .. 367 .. 370 .. 374 .. 375 .. 4/04/84: 375.2 . 382.2 HFB MUCH HFB HFB MUCH MUCH MUCH 3 MUCH 3 MUCH 3 MUCH 3 MUCH 3 MUCH MUCH MUCH Main breaker on HFB tripped out. Problem was a poor HFB connection on load side of the breaker. Breaker was overheated in that area. Connection was tightened, resolving the tripping problem. Breakers on intermediate vehicles 6 and 7 tripped out, reason unknown. Breakers were reset. HFB conveyor hydraulic motor hose failed. The crimp failed, allowing hose to blow out end of the fitting. Crimp-type staple-lock fitting was replaced. HFB was jammed. Wet coal overloaded conveyor system. Hopper was emptied manually. Also, 1 hydraulic fitting on conveyor motor was leaking its O-ring was replaced. Breaker on intermediate vehicle 1 tripped out, reason unknown. Discharge vehicle sheared a pin during system troubleshooting. Breaker was reset and shearpin was replaced. Breaker on intermediate vehicle 4 tripped out, reason unknown. Breaker was reset. Intermediate vehicle 3 conveyor drive chain broke and taper lock bushing bolt (on conveyor drive sprocket) sheared. Motor fan was also hitting against guard. Chain was repaired; bolts and fan were replaced. System shut down for unknown reason. System started up without a problem. Intermediate vehicles 7 and 8 breakers tripped out, reason unknown. Breakers were reset. Intermediate vehicles 4 and 6 breakers tripped out, reason unknown. Breakers were reset. Intermediate vehicles 4 and 6 breakers tripped out, reason unknown. Breakers were reset. Intermediate vehicles 4, 6, and 7 breakers tripped out, reason unknown. Breakers were reset. Intermediate vehicles 1, 2, 3, and lead vehicle did not start. Problem was traced to a loose wire in electrical XP box on intermediate vehicle 4 (control power circuit). Instantaneous overload in breaker was set to its highest level, eliminating the nuisance tripping problem in intermediate vehicle 4. Breaker on intermediate vehicle 3 tripped out, reason unknown. Breaker was reset. See explanatory notes at end of tabulation. 54 Cumulative run time, min Av loading rate, st/min Machine 4/04/84: 402.2 1.5 1.5 1.5 1 .8 .8 .8 3 3 3 3 4 4 5 5 MUCH 422.2 MUCH 436.2 MUCH 4/05/84: 439.2 MUCH 443.2 MUCH 446.2 MUCH 447.2 MUCH 517.2 MUCH 04/06/84: 518.2 HFB 525.2 MUCH 525.7 HFB 531.7 HFB/MUCH MUCH 536.7 4/09/84: 537.2 MUCH 548.2 MUCH 550.2 MUCH Description Intermediate vehicle 3 sheared a pin. Pin was replaced. Clutch on intermediate vehicle 1 slipped. It was reset. Clutch on intermediate vehicle 1 slipped; it could not be reset. Clutch was taken out and replaced with a solid coupling (Lovejoy). Intermediate vehicle 6 sheared a pin. Shearpin was replaced. Clutch on intermediate vehicle 10 slipped. It was reset. Breaker on intermediate vheicle 1 tripped out, reason unknown. Breaker was reset. Breaker on intermediate vehicle 1 tripped out, reason unknown. Breaker was reset. Breakers on intermediate vehicle 5 and the discharge vehicle tripped out (thermal overload). After discharge vehicle motor cooled, breakers were successfully reset. HFB was jammed. Ran conveyor back and forth to clear hopper. Intermediate vehicle 1 stopped. Reason unknown. HFB was jammed. Manually shoveled the system and ran conveyor back and forth to clear system. HFB was jammed. Manually shoveled the system and ran conveyor back and forth to clear system. Intermediate vehicle 1 also stopped, reason unknown. Clutch on intermediate vehicle 10 slipped; it would not reset. Clutch was removed and replaced with a solid coupling (Lovejoy). Intermediate vehicle 3 stalled. It appeared that the overload trip was caused by fines buildup. Breaker on intermediate vehicle 8 tripped out, reason unknown. Breaker was reset. Intermediate vehicle 10 conveyor drive gearbox overheated. Temperature after 15 min of operation was 239° F. Haulage rate was 5 st/min average. It was found that the fines buildup around the conveyor drive chain had caused it to seize, significantly increasing the system load, and thereby increasing the gearbox temperature. After end of run on April 9, chain was removed, cleaned, and reinstalled. Gearbox no longer overheated. Breaker on intermediate vehicle 8 tripped out, reason unknown. Breaker was reset. See explanatory notes at end of tabulation. 55 Cumulative run time, min 4/09/84: 550.7 551.7 552.2 4/12/84: 553.2 554.2 556.2 556.3 556.4 556.5 556.6 560.6 Av loading rate, st/min Machine Description 5 MUCH 6 MUCH 6 HFB 5 MUCH 4 HFB 4 MUCH 4 MUCH 4 MUCH 4 MUCH 4 MUCH 4 MUCH Breaker on intermediate vehicle 8 tripped out, reason unknown. Breaker was reset. Breaker on intermediate vehicle 8 tripped out, reason unknown. Breaker was reset. HFB conveyor hydraulic motor hose blew out of crimp fitting. (This hose was on the opposite port to hose that failed on 3/30/84.) Hose was replaced. Lead vehicle sheared pin. Shearpin bushing was also damaged. Pin and bushing were replaced. HFB was jammed. Ran conveyor back and forth to clear system. Breakers on intermediate vehicles 7 and 8 tripped out, reason unknown. Breakers were reset. Breakers on intermediate vehicles 7 and 8 tripped out, reason unknown. Breakers were reset. Breaker on intermediate vehicle 7 tripped out, reason unknown. Breaker was reset. Breaker on intermediate vehicle 8 tripped out, reason unknown. Breaker was reset. Breakers on intermediate vehicles 7 and 8 tripped out, reason unknown. Breakers were reset. Breakers on discharge vehicle and intermediate vehicle 10 tripped out. Wire on load side of main breaker was loose. Connections were tightened and breakers were reset. HFB Hopper-feeder-bolter. MUCH Multiple-unit continuous haulage system. XP Explosion-proof. 56 APPENDIX C.-MUCH SYSTEM COAL CARRYBACK LOSS-CONVEYOR TEST- JULY 31 TO AUGUST 3, 1984 Weight of coal, lb Lead vehicle: Outby end 369 Intermediate vehicle 1: Conveyor drive shaft area 215 Intermediate vehicle 2: Conveyor drive shaft area 318 Intermediate vehicle 3 (modified), 2,504 lb lost: Hopper cleanout port 225 Conveyor drive shaft area and pelican beak 681 Midship cleanout port 780 Outby cleanout port and intermediate vehicle 4 hopper cleanout port 818 Intermediate vehicle 4 (modified), 1,853 lb lost: Conveyor drive shaft area and pelican beak 353 Midship cleanout port and chain slots 668 Rear chain slots and outby cleanout port 832 Intermediate vehicle 5: Conveyor drive shaft area 155 Intermediate vehicle 6: Conveyor drive shaft area (caps open) 147 Intermediate vehicle 7: Conveyor drive shaft area 617 Intermediate vehicle 8 (modified), 3,240 lb lost: Hopper cleanout port 253 Conveyor drive shaft area 535 Midship cleanout port, chain slot, and outby 2,632 Intermediate vehicle 9 (vehicle misaligned) 109 Average loss of unmodified vehicles 264 Average loss of modified vehicles 2,258 Increase in coal spillage 854 57 APPENDIX D.-CONVEYOR ACCEPTANCE TEST LOG-OCTOBER 29, 1986 Clock Run time, min Delay time, min time Run Cumulative Delay Cumulative a.m.: 8:00 8:20 20 20 8:23 3 23 8:25 23 2 2 8:27 2 25 2 9:07 40 65 2 9:08 65 1 3 9:10 2 67 3 9:37 27 94 3 9:38 94 1 4 9:45 7 101 4 9:47 2 103 4 9:50 3 106 4 9:52 106 2 6 9:55 3 109 6 9:57 109 2 8 10:00 3 112 8 10:02 112 2 10 10:05 3 115 10 10:06 115 1 11 10:16 10 125 11 10:17 125 1 12 10:26 9 134 12 10:28 134 2 14 10:30 134 2 16 10:42 12 146 16 10:45 146 3 19 11:05 146 20 39 p.m.: 12:20 75 221 39 12:30 221 10 49 2:00 90 311 49 2:15 311 15 64 2:30 15 326 64 2:34 326 4 68 2:45 11 337 68 NOTE.-Total test time-405 min, total run time-337 min, total down time-68 min. Event Start test, 6 st coal on system. 3 st coal to system. System shutdown, lead vehicle conveyor motor. Restart system. 3 st coal added to system. System shutdown. Restart system. 2 st coal to system. System shutdown. Restart system. 1 st rock to system. 1 st rock to system. System shutdown, lead vehicle conveyor motor. Restart system, add 1 st rock. System shutdown, intermediate vehicle 2, conveyor motor. Restart system. System shutdown, intermediate vehicle 5 conveyor motor. Restart system. System shutdown, lead vehicle conveyor motor. Restart system. System shutdown, lead vehicle conveyor motor. Restart system. System shutdown, lead vehicle conveyor motor. Removed 2 st material from system. Restart system. System shutdown, lead vehicle conveyor motor. Adjusted thermal trip setting on lead vehicle. Restart system. Stop system, repair coupling on conveyor drive motor discharge vehicle. Restart system. System shutdown, lead vehicle conveyor motor. Restart system. System shutdown, intermediate vehicle 5 conveyor motor. Restart system. Stop test. 58 APPENDIX E.-RECOMMENDED SPARE PARTS FOR IN-MINE TRIAL Item Quantity 7.5-hp tram motor, frame 215 TC 1 5.0-hp tram motor, frame 215 TC 2 Cone drive tram speed reducer 1 1.0-hp hydraulic pump motor, frame 182 T 1 Lovejoy couplings, 1-1/2-in x 5/8 in 6 Speed switch chain, Morse 40 5B 80 pitch 4 Speed switch sprocket 4 Conveyor drive sprocket, 14-tooth 4 15-hp conveyor motor, frame 254 TDZ 2 Conveyor drive shaft with housing and bearings 2 Conveyor toe shaft assembly 2 Conveyor tail shaft assembly 2 Tram brake assembly 1 Headlight assembly 2 Conveyor speed reducer, lead vehicle 1 Conveyor speed reducer, intermediate and discharge vehicles 2 Tires on rims 2 Drive axle assembly (1 intermediate and 1 lead) 2 Nondrive axle assembly 2 Conveyor chain, including flights 2 59 APPENDIX F.-MUCH SYSTEM MODIFICATIONS SUMMARY Reason Modification Voltage of electrical control circuits were reduced from 460 V ac to 120 V ac line-to-line. 460- V primary to 120- V secondary, 2-kV»A stepdown control transformer installed in the electrical control circuit. 15-A fuses added to control circuits. Main circuit breaker (CB-1) for system, shunt trip coil, and all magnetic coils for motor control contactors and time-delay relays were replaced. Emergency stop electrical circuit modified to provide fail- safe-operation. Size of system trailing cable increased from a 1/0, 3-con- ductor, 90° C rated cable to 3/0, 3-conductor, 90° C rated cable. 1/0 size cable between the discharge and intermediate vehicle 10 replaced with 3/0 cable, and 1/0 cable be- tween intermediate vehicles 10 and 9 and beteen 9 and 8 replaced with 2/0 size power cable. Electrical fault ground-check circuit was originally connected to grounding studs in the explosion-proof (XP) box of each vehicle. This was modified so that circuit was only connected to ground in the lead vehicle XP box. Electrical connection box for the discharge vehicle headlight circuit replaced with 2-pole lever action pushbutton switch mounted in XP enclosure. All electrical motor thermal overload relay heater elements were replaced. Lead vehicle operator canopy was redesigned and load tested. Frame structure of each vehicle was modified to elevate discharge end of conveyor deck by 5 in. Cable support trays were fabricated and installed through the drawbars between vehicles to support cables and water hose between vehicles. Discharge vehicle steering system was modified by fabricating a new bell crank arrangement and relocating steering cylinder. MSHA memorandum stated "The voltage of alternating current control circuits shall not exceed nominal 120 V line-to- line." Transformer was required to step down control circuit voltage from 460 V ac to 120 V ac. To protect the control circuit components from overcurrents. Changes required because of reduction of control circuit voltage from 460 V to 120 V. Safety. Amperage of 1/0 cable was insufficient to handle normal- ly expected amperage requirements of system. Do. Original grounding was not MSHA approvable because an open ground fault in an individual vehicle would not be detected. Connection box did not meet MSHA specifications. Original overload elements were improperly sized to adequately protect the motor. Original canopy was damaged during testing. To reduce vehicle-to-vehicle interference during tramming over uneven terrain. To help reduce cable and hose damage between vehicles. More steering angle was needed to improve steering ability of discharge vehicle. 60 Modification Pull-cord type emergency shutdown system was installed for surface testing of the MUCH. Conveyor chain slack adjustment system was designed and installed on vehicles. Conveyor chain holddowns were installed on vehicles. Conveyor deck extension plate was added to each vehicle to bridge gap between end of the conveyor deck and conveyor chain tail shaft. Shearpins located in conveyor drive mechanism were removed and shearpin couplings were welded solid. Autoguard torque clutches were removed from conveyor drive system and replaced with Lovejoy couplings. Reason To provide complete shutdown of entire system from either side along the length of the system. To reduce amount of fines buildup on conveyor decks. To prevent fines buildup on conveyor decks. To prevent material from getting jammed between end of conveyor deck and tail shaft. During conveyor testing, the shearpins proved to be unreliable. Clutches failed and could not be reset. * U.S. GOVERNMENT PRINTING OFFICE: 611-012/00,124 INT.BU.OF MINES.PGH..PA 28977 N> 00 c m g> o5 c/> z «L -* <» • > 5 ■n O 33 o ■n O > Departm au of Mil E Street iiington, ■o r~ a- £ the Int , MS# 20241 W z m O c > r - O -o -o O 30 -< m "0 r - O -< m DO 105 90 ? • A^^ «• y *+ *by Ap° *<£►**•' .r>~ . t • o„ *!> V ♦ ^ A^ ^*y .v^BV. ^v A « *£M>s. *>„& 6°* r + V< ^ aP 'I,*-' ^ v •••• S.J' :'Mk. \S :' 6°*. 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